United States Patent [19]
Reynolds et al.
[11] Patent Number:
[45] Date of Patent:
4,556,422
* Dec. 3, 1985
[54] PROCESS FOR THE RECOVERY OF LEAD
AND SILVER CHLORIDES
[75] Inventors: James E. Reynolds, Golden; Alan R.
Williams, Denver, both of Colo.
[73] Assignee: Hazen Research, Inc., Golden, Colo.
[ * ] Notice: The portion of the term of this patent
subsequent to Jun. 30, 1998 has been
disclaimed.
[21] Appl. No.: 255,649
[22] Filed: Apr. 20, 1981
1,951,342 3/1934 Bradley 75/77
3,172,753 3/1965 Walsh 75/77
3,764,490 10/1973 Chambers 75/101 R
4,082,629 4/1978 Milner , 423/98
4,135,993 1/1979 Um 423/98
4,276,084 6/1981 Reynolds 423/98
OTHER PUBLICATIONS
International Critical Tables, vol. 7, (1st ed.),
McGraw-Hill, N.Y., 1930, pp. 313-316.
Primary Examiner-John Doll
Assistant Examiner-Robert L. Stoll
Attorney, Agent, or Firm-Sheridan, Ross & McIntosh
27 Claims, No Drawings
A process for selectively leaching lead and silver chlorides
from a sulfide ore residue in a rapid time which
comprises brine leaching the residue under pressure at a
temperature above the normal boiling point of the solution,
preferably above 100° C.
Modifications are leaching at the agglomeration temperature
of sulfur when present in the residue to agglomerate
the sulfur for ease of recovery, and flashing
from leach temperature to ambient as a lead chloride
crystallization recovery step to produce a large crop of
lead chloride crystals per pass.
Related U.S. Application Data
[63] Continuation-in-part of Ser. No. 80,444, Oct. I, 1979,
Pat. No. 4,276,084.
[51] Int. Cl.4 C22B 13/04; COlG 21116;
COlG 5/02
[52] U.S. Cl 75/101 R; 75/114;
75/120; 75/118 R; 75/78; 423/27; 423/39;
423/94; 423/98
[58] Field of Search 423/27, 39, 98, 94;
75/118 R, 120, 77, 78, 101 R, 114
[56] References Cited
U.S. PATENT DOCUMENTS
1,396,740 11/1921 Ganelin , 423/98
[57] ABSTRACT
4,556,422
1
PROCESS FOR THE RECOVERY OF LEAD AND
SILVER CHLORIDES
RELATED APPLICATIONS
This application is a continuation-in-part of my U.S.
patent application flIed Oct. 1, 1979, Ser. No. 080,444,
U.S. Pat. No. 4,276,084 entitled "Hydrometallurgical
Process to Recover Lead from Lead Concentrates."
TECHNICAL FIELD
The invention lies in the field of recovery of the
chlorides of lead and silver by selectively solubilizing
the chlorides from other solid materials and the final
recovery of the metals from the solubilized chlorides.
BACKGROUND ART
2
silver chlorides, it has been found effective for selectively
leaching lead and silver chlorides at a rapid rate
from residues obtained by cupric chloride leaching of
pyritic or sulfidic lead ores in which lead and silver
5 chlorides are selectively leached at saturation into solid
products, all as disclosed in my continuation-in-part
application referred to above. Among other materials,
these residues contain the sulfides of copper, iron and
zinc. The process is equally effective for selectively
10 leaching the residues produced from halogen leach of
pyritic ores in the processes of U.S. Pat. Nos. 4,135,993
and 4,173,623. The operation of the invention will be
illustrated by its application for the recovery of lead
from the residue resulting from the cupric chloride
15 leach of sulfidic ores described in the above-referred-to
continuation-in part application.
Pressure Brine Leach
The solids from the cupric chloride leach reaction,
comprising lead chloride, silver chloride, elemental
sulfur, unreacted metal sulfides of other metals, and
gangue, are treated for the selective separation of lead
and silver chlorides. The lead and silver chlorides are
selectively solubilized from the residue in the illustrative
embodiment by leaching with an aqueous brine
solution having a sodium chloride concentration of
from about 200 gpl to saturation. Suitable substitues for
sodium chloride are the other alkali metal chlorides,
lithium and potassium chlorides, as well as the alkaline
earth metal chlorides, calcium and magnesium chlorides.
When other solutes than sodium chloride are used
the upper limit of the amount used will change, the the
minimum amount of solute preferably being above
about 200 gpl. A leach pH of about 0-7 is preferred. Use
of too high a pH will precipitate lead compounds. Hydrochloric
acid (hydrogen chloride) may also be used as
one of the chlorides. The solute must be a chloride
which provides maximum chloride ion concentration to
40 the saturation point under the reaction conditions. The
solubilization of silver chloride may be enhanced by the
use of an oxidant in the leach, such as, sodium chlorate
or oxygen.
The brine leach is conducted at a temperature in
excess of the solution boiling temperature, which, of
course, requires a pressurized system. The temperature
is maintained between about 100· C. to about 170· C.,
the system pressure being selected so as to accommodate
the solution temperature while preventing solution
boiling. Pressures from about 30 to about 150 psig are
suitable to accomplish this purpose. If elmental sulfur is
present in the residue the agglomeration temperature of
sulfur is used. This is about 130· C. to 140· C. It was
found that when the sulfur is agglomerated during the
leach and separated from the liquid chlorides in this
form with the other solids it can be readily separated
from the other solids by physical methods, such as, wet
screening.
The brine leach, conducted under the described temperatures
and pressures, accomplishes a relatively high
solubility of lead and silver chlorides in a relatively
short period of time, while leaving the elemental sulfur
and unreacted metal sulfides in the residue phase. Retention
times of from about 30 seconds to about 5 min-
65 utes are generally adequate to dissolve lead chloride to
solution concentrations of at least about 130 grams per
liter of lead. A preferred leach time is not in excess of
about three minutes. Increased lead concentrations as a
DISCLOSURE OF THE INVENTION
Prior Art Statement
U.S. Pat. No. 4,113,471 discloses a process for brine 20
leaching oxide ores to non-selectively solubilize nonferrous
metal values as chlorides, at elevated temperatures
and pressures with the addition of oxygen. Leaching
is required for a time of ! to 12 hours. There is no
:~~~~~v~r~~~ching of lead or silver chlorides from their 25
U.S. Pat. Nos. 4,135,993 and 4,173,623 teach brine
leaching lead chloride at temperatures of 80·-120· C. to
selectively solubilize the lead chloride out of a sulfidic
residue containing the sulfides of copper, iron and zinc
as well as elemental sulfur. The leaching is not done 30
under pressure. Leach times of ! to 2 hours are required
in both processes.
International publication No. WO 80/00852, class 22
B 13/00, published under the patent cooperation treaty
on May 1, 1980 (01.05.80), discloses the recovery oflead 35
from crystallized lead chloride by reduction with hydrogen
or a hydrogen containing compound accompanied
by condensing lead chloride volatilized during the
reduction process.
BEST MODE FOR CARRYING OUT THE
INVENTION
Although the process is not limited in its application
to any particular starting materials containing lead and
Lead and silver chlorides are selectively separated
from other solid materials by leaching the materials
with a brine leach at elevated temperatures and pressures
to selectively solubilize the chlorides, followed by 45
a liquid-solids separation. A starting material on which
the process is particularly effective is the sulfidic resi- .
due obtained by selectively leaching a sulfidic lead ore
with a cupric or a halogenating leach to produce solid
lead chloride. Temperatures ranging from the boiling 50
point of the solution to 170· C. are used to selectively
solubilize substantially all of the lead and silver chlorides
in a rapid time which can be not in excess of about
three minutes. If elemental sulfur is present in the starting
material, it is agglomerated at elevated temperatures 55
for ease in separation from the solid residue. Lead chloride
is crystallized from solution by flashing from the
high temperature of the leach to lower temperatures to
produce a large crop of lead chloride crystals per pass
and elemental lead of high purity recovered from the 60
crystallized lead chloride by hydrogen reduction, or
otherwise. Silver chloride is recovered from the mother
liquor and processed for recovery of silver.
3
4,556,422
4
89
93
92
Pb
Wt%Pb Extraction (%)
26.1 gil
20.6%
28.6 gil
18.7%
28.5 gil
17.2%
Vol.
120ml
1.9 g
IlOml
I.3g
625 ml
9.1 g
TABLE 2-continued
PF
Residue
PF
Residue
PF
Residue
3
Distribution
Size Weight S' Weight S'
Fraction g % % %
Plus 200 (beads) 1.67 81.8 33.9 89.5
Minus 200 (fines) 3.26 ~ ~ ~
Totalloverall 4.93 (31.0) 100.0 100.0
10
Rate of Brine Leaching Test
galena in L14 liter of 90 gil Cu+2 and 200 gil
NaCI at pH = I (HCI) for one hour
at 60' C.
Brine Leach Solution: I liter 250 gil NaCl, pH 1.5.
Procedure: Feed added to solution with continuous
stirring. Thief samples removed at
designated time and immediately vacuum filtereg
without rinsing. .
Leach
Time
(min.)
40 5.0 g, S' - agglomerated product from Test 1151-107-1:
Conditions: Test 1151-99-1 leach residue
130' C.
pH 11.8 with KOH
2 Hours
Elemental Sulfur Distribution in ± 2oo-mesh Size
Fractions of S'-Agglomerated Autoclave Leach Residue
Feed to Wet Screening
45
Tem- Sample Diluted Diluted Total Pb
per- Vol- to Sample Pbin Solubilature
ume Volume Pb Sample ity
'C. ml ml gil g gil
42 10.0 200 0.746 0.149 14.9
70 10.0 200 1.45 0.290 29.0
94 10.0 200 2.94 0.588 58.8
1001 87.5 795 9.41 7.48 85.5
1241 76.0 1360 6.20 8.43 110.9
1441 79.0 1945 7.13 13.87 175.5
47.5 10.0 200 0.972 0.194 19.4
69.0 10.0 200 2.97 0.594 59.4
96 10.0 200 6.02 1.204 120.4
1001 8.5 1320 8.96 11.8 139.1
1241 8.3 2060 6.67 13.7 165.5
1352 9.2 1625 11.0 17.9 194.3
1431 8.2 2200 8.82 19.4 236.6
240
320
NaCI
gil
result of high temperature and pressure brine leach
significantly facilitate further separation. processing.
The brine leach may be conducted at lower pressures,
including atmospheric pressure, and lower temperatures,
as in the prior art. However, pressures and tern- 5
peratures lower than those recited for the preferred
range of the brine leach will require more of the brine
solution per amount of lead and silver chlorides and a
longer retention time in order to solubilize the chlorides.
10
In a·typical application, washed tails or residue as. a
filter cake. from the above-referred-to cupric leach of
sulfidic ores was brine leached in an externally heated
concentric double pipe pressure leach. Brine containing
280 gil of NaCl lmd PbC1zcake was heated to 135° C. 15
under 50 psig pressure with a one minute retention time
to dissolve PbC12UP to a concentration of 145 gil Pb.
This procedure reduces the size of the crystallizer used
in subsequent.PbCh crystallization and circuit •flows to
a fraction of that of an ambient pressure system with a 20 Sulfur Agglomeration
corresponding reduction in heating and cooling needs. The agglomeration of sulfur is accomplished during
To explore the •effectiveness .of. the brine pressure leaching. by operating the brine leach within the sulfur
leach at high temperatures in rapidly solubilizinglarge agglomeration temperature range thereby permitting
amounts of lead chloride, the solubility system PbCI2- the sulfur to be readily separated from the remainder of
NaCI-H20 was extended to 144°C. at two hrinecon· 25 the residue following liquid-solids separation. Agglomcentrations
of 240 and 320 gil of NaCI used for leaching eration tests were run on a brine leach residue from
a lead sulfide ore residue from a cupric leach, the com- cupric. leach of lead sulfide ore as referred to above
position ofthe residue being typified by the brine leach containing elemental sulfur to see if the sulfur could be
residues of Examples 1.and 2. The results of the solubil- coalesced to a size large enough for a wet screen separaity
tests reported in Table 1 below show a decided 30 tion. The autoclave leach was made at 130° C. The
nonlinear increase in the solubility oflead chloride with results recorded. in Table 3 below shows that the plus
increase in temperature above the. boiling point. of the 200 mesh fraction contains about 90 percent•ofthe free
solution,and particularly above 125°C. sulfur with a grade of 82 percent, thus shoWing that the
TABLE 1 35 procedure is feasible for sulfur separation.
------,-:.;:...:::.::.::...:-------- TABLE 3
'Sample withdrawn from pressurized autoclave (150 psig N2) using sample bomb. 50
2Repeat run to check data, new solutions.
Since a pipeline brine leach is contemplated in the
most feasible commercial application of the process, a
minimum leach time is required in the interest of reducing
equipment cost and processing time. Rate of brine
leaching tests at high temperatures were made on a
residue obtained by the above-referred-to cupric leach
of a lead sulfide ore. The brine leach contained 250 gil
of NaC1 and a pH of about 1.5 was used. The leach
temperature was 140° C. The results recorded in Table
2 below indicate that substantially all of the PbClz is
leached in a time not in excess of about three minutes.
Liquids-Solids Separation
Following the brine leach, the pregnant lead and
silver chloride solution is separated from the remaining
55 residue for subsequent recovery therefrom of lead and
silver values. As high temperatures and pressures are
utilized during the leach, the liquid-solid separation
must be conducted under pressure in order to prevent
flash crystallization of the lead chloride from the solu-
60 tion. One suitable technique to accomplish the separation
while avoiding flash crystallization is to employ
small diameter pressurized liquid cyclones in parallel,
the hydroclones operating to permit pressure reduction
TABLE 2 to atmospheric as the cyclone operation effects a liquid-
--------....::~::..::=--=---------65 solids separation. A pressure drop of about 40 psi across
Rate of Brine Leaching Test the cyclone system occurs. Hydroclone techniques such
Feed: 50 g cu+
2
grade galena contlaeiancihnegd5r8e.6si8dupeerocfenathPigbhand· as those discussed in The Hydroclone, D. Bradley, Perobtained
by leaching 200 g of a high grade gamon Press, Lrd. 1965 may be utilized in this context.
4,556,422
6
material which is substantially impervious to the corrosive
action of lead chloride, such as, castable or refractory
brick. The materials of which the reactor walls
must be made have such a low heat conductivity that it
is practically impossible to heat the reactor contents
with heat applied to the outside of the walls. Accordingly,
it was necessary to devise a practical procedure
for internally heating the reactor contents to a temperature
up to 900· C. at least. Two alternative procedures
were found to be feasible.
In accordance with one procedure a furnace or reactor
made of refractory brick was used. Heat for the
endothermic reaction occurring in the reaction chamber
was supplied by fire tubes submerged in molten lead
in contact with lead chloride and the other reactants in
the reaction chamber. Means are provided for introducing
reactants into the reaction chamber and for continuously
or intermittently tapping pure lead from the furnace.
Means are also provided for condensing vaporized
lead chloride and returning the vaporized lead
chloride to the reaction chamber. Lead chloride does
not react with molten lead and having a lesser specific
gravity floats on top of the molten lead.
The second procedure comprises introducing into the
reaction chamber a partially uncombusted gas mixture
supplying hydrogen, and completing the combustion
with oxygen in an endothermic reaction which supplies
heat for the endothermic lead chloride reduction reaction.
Heat balance calculations showed that sufficient
30 heat can be brought into the lead chloride reduction
reactor to supply the endothermic heat of reaction and
other heat requirements, including that caused by heat
loss, by using a reducing combustion gas or gas mixture,
the term "gas" as used herein and in the claims including
both. Any hydrocarbon or mixture of hydrocarbons
which supply hydrogen can be used. A mixture produced
by a partial combustion ofhydrocarbons, such as,
methane or propane, provides both the hydrogen and
the heat needed for the endothermic reduction of lead
chloride. Contrary to what might be expected, introduction
into the reaction area of large volumes of water
vapor formed in the partial combustion reaction and
diluent gases does not adversely affect the reduction
reaction. The above described procedure applies also to
the recovery ofcopper from cuprous chloride by reduction
of hydrogen.
Illustrative gases and gas mixtures found suitable are
H2-CO-C02, H2-CO-C02-H20, and H2-CO-
N2. The gases used mayor may not be supplemented
by hot reducing combustion gas. Oxygen gas or
air may be used to supply oxygen.
The PbCh cake was metered to a brick-lined PbCh
reduction furnace as described above operating between
600·-900· C., preferably at about 800· C. A reducing
gas feed of 98 percent H2 from an on-site H2
plant was used. An excess of 240 percent of theoretical
H2 was fed based on lab tests in batch tube furnace runs.
This produces an exit gas consisting of 60 percent HCI
and 40 percent H2, by volume. Some volatilized PbCh
leaves the reactor zone with the off-gas but is refluxed
back to the furnace by either a molten lead splash condenser
or an air-cooled surface condenser. Any additional
heat requirements for the endothermic reduction
reaction and to bring reactants up to temperature may
be supplied by indirect firing of submerged fire tubes in
the molten lead in the reactor as described above. High
purity lead is tapped continuously or intermittently
from the furnace into a casting machine.
5
Another solids separation device, such as, an insulated
or jacketed pressure leaf filter can be used to accomplish
the same objective.
In operation, the pipeline dissolver discharges
through a bank of 10 mm alumina cyclones to remove 5
solids at about a 4-5 micron cut point with a let down
from a 50 psig pipeline leach to atmospheric, the pressure
being utilized to remove the solids. Flocculant may
be injected at the cyclone inlet to improve clarity of the
cyclone vortex flow. The apex flow, containing unre- 10
acted sulfides and agglomerated sulfur, flashes to atmosheric
pressure and mixes with concentrate and
mother liquor from the subsequent PbCh crystallization
to quench the hot slurry and solidify beads of agglomerated
sulfur. The slurry is gravity-fed to a wet screen or 15
similar separation device to make a separation of agglomerated
sulfur beads from other solids, principally,
unreacted sulfide tails. The fines are dewatered and
fmally filtered by conventional filtration. Filtrate is
recycled to the leach feed tank and tails cake is dis- 20
charged to a solids disposal area. Prior to reaching the
leaching tank the leach can be purified by a bleed
stream in which copper and lead values are recovered
by iron cementation and soda ash used at pH 9 to precipitate
Fe, Mg, and Zn to permit recycle of barren 25
brine. The residue from sulfur separation is disposed of
or further processed for recovery of metal values if
warranted.
Lead Chloride Crystallization
Lead chloride is crystallized from the liquid phase
resulting from the liquid-solids separation on the brine
leach solution for subsequent recovery of elemental
lead by hydrogen reduction, or otherwise. Two-stage
crystallization may be used with the first-stage at atmo- 35
spheric pressure and the second stage at about 50 mm
Hg absolute to cool the feed to about 40· C. A pregnant
brine containing up to 145 gil Pb flashes typically from
135· C. to ambient temperature in the first stage to
produce a large crop of crystals per pass. Surface con- 40
densers may be used for the second stage, with contaminated
lead chloride condensate being recycled to process.
Mother liquor overflow and crystal withdrawal
elution leg are specific design requirements to elute
minus 5-micron impurities not removed in the cyclones. 45
Alternatively, polish fIltration techniques could be used
to separate minus 5-micron solids.
Crystallizer under-flow is removed through an elution
leg at 40-50 percent solids and advanced to a washing
centrifuge. A three percent moisture PbCh cake is 50
conveyed to a surge hopper above the PbCh reduction
furnace.
Lead Recovery from Lead Chloride
The lead chloride is reduced to high purity lead by 55
hydrogen directly without further refining. The remaining
solubilized silver chloride is treated for recovery of
silver by cementation or other means. Other conventional
methods may be used to recover elemental lead
from the lead chloride. Hydrogen supplying com- 60
pounds, such as, methane and propane may be used as a
source for hydrogen.
Since the reduction of PbCh is endothermic, heat
must be supplied to the reaction represented by the
formula PbCh+H2Pb+2 HCI when an excess of hy- 65
drogen over stoichiometric is used. As lead chloride is
extremely corrosive, the reactor cannot be made of
conventional reactor materials but must be made of
Sb
2.0 0.33
Zn Fe
Brine Leach Residue Assay, %
Ag
Cu (oz/ton)
We claim:
1. A process for solubilizing a chloride selected from
the group consisting of lead chloride and a combination
of silver and lead chlorides which comprises subjecting
the chloride to a brine leach at a temperature above the
normal boiling point under a pressure above atmospheric
sufficient to prevent boiling for a sufficient time
to solubilize substantially all of the chloride.
2. The process of claim 1 in which the temperature is
maintained above the normal boiling point to about 1700
C.
3. The process of claim 1 in which the time of leach
8
sufficient hydrochloric acid to maintain a pH of about 1.
The cupric chloride leach was conducted at a temperature
of 600 C. for two hours. The residue of the cupric
chloride leach was subjected to a 900 milliliter brine
5 leach at a temperature of 800 _900 C. and about atmospheric
pressure of one-half hour. The brine solution
contained about 250 grams of sodium chloride per liter.
The analysis of the brine leach residue, which weighed
83 grams, and the results of the extraction are set forth
10 in Table 5.
The extraction resulted in 19.0 grams oflead chloride
being produced. This lead chloride was reduced to lead
in an atmosphere of 175 cubic centimeters per minute of
hydrogen, 75 cubic centimeters per minute of carbon
15 monoxide, 75 cubic centimeters per minute of carbon
dioxide at a temperature of 8000 C. for 35 minutes. The
lead metal was assayed by emission spectroscopy. The
lead metal was 99.98 percent pure. It contained impurities
of 0.01 percent silicon, 0.005 percent iron, 0.001
percent copper and 0.001 percent bismuth with no other
elements being detected.
TABLE 5
4,556,422
EXAMPLE I
7
Off-gas is scrubbed in a packed tower or similar
scrubbing device using liquor from cupric leach, and a
large excess of dilution air to lower H2 content to a safe
level and also simultaneously consume scrubbed HCl
which may be used to reoxidize the cuprous ion to
cupric. A water scrubber may also be used to recover
the HCI. Exit gas, free of HCI and particulate matter, is
exhausted to atmosphere.
Up to over 99 percent of lead was obtained from the
starting material. Lead having a purity of +99.9 percent
was consistently obtained by the process. The
recoveries of lead and silver shown in Tables 4 and 5, as
produced by Examples 1 and 2, are representative of
recoveries obtained by the process. The lead purity
obtained in Example 2 is also typical.
Two different 100 gram samples of a lead concentrate
having a composition of 18 percent lead, 26.2 percent
zinc, 0.54 percent copper, 5.1 troy ounces of silver ore 20
ton of concentrate, 0.029 percent antimony and 14.4
percent iron were treated with 250 milliliters of a cupric
chloride leach solution comprising about 50 grams of
copper per liter as cupric chloride and 200 grams of
sodium chloride per liter. The pH of the leach solution 25 Pb
was maintained at about 1 through the addition of hydrochloric
acid. After 3 hours, a total of 4.08 and 4.80 0.21 0.805 Ex~:~~tion. 2~8
grams ofhydrochloric acid were added to Sample 1 and Pb Cu Zn Fe Ag Sb .
Sample 2, respectively. The cupric chloride leach of
Sample 1 was conducted at a temperature of 600 C. and 30 __9_9_.4 -_4..,;8;,...5 3._6__7_._1 8_1._6 ;.;32;...-_
the cupric chloride leach of Sample 2 was conducted at
a temperature of 800 C. The residue of the cupric chloride
leach of each of the samples was separately brine
leached in a brine solution containing about 250 grams
of sodium chloride per liter at a temperature of 800 _850 35
C. and about one atmosphere for one-half hour. Each
brine leach slurry was filtered while hot and the residue
was washed fIrst with hot brine solution and then with
water. The analyses of the brine leach residue and the
results of this extraction are set forth in Table 4. The 40
negative extracted copper percentages are due to a
portion of the cupric chloride of the leach solution
being precipitated to copper sulfIde.
TABLE 4
Cupric
Leach Brine Leach Residue Assay, %
Time Weight Ag Extraction, %
(Hours) Product (gm) Pb ZN Cu FE (oz/ton) Sb Pb Zn Cu Fe Ag Sb
Sample 1:
1.0 1hr. 4.32 0.49 36.1 0.78 20.3 1.4 0.013 98.3 11.8 7.9 7.6 82.4 71
residue
3.0 Final 50.7 0.11 36.1 0.96 20.0 1.8 0.014 99.6 18.7 -4.5 18.1 79.2 72
residue
Sample 2:
1.0 1hr. 4.43 0.10 37.4 0.98 19.8 1.2 0.Q18 99.6 7.2 -17.5 10.6 84.7 60
residue
2.0 2 hr. 4.52 0.08 36.7 1.00 19.2 1.1 0.Q11 99.7 17.4 -8.9 21.3 87.3 78
residue
3.0 Final 48.3 0.06 36.2 1.44 19.8 1.2 0.Q15 99.8 22.6 -49 23.0 86.8 71
residue
EXAMPLE 2
A 125 gram sample of a lead concentrate having a
composition of 25.1 percent lead, 9.57 percent zinc, 0.36
percent copper and 16.3 percent iron was treated with
500 milliliters of a cupric chloride leach solution comprising
about 50 grams of copper per liter as cupric
chloride, 200 grams of sodium chloride per liter and
is not in excess of about three minutes and a concentration
of at least about 130 gpl oflead chloride is obtained.
4. The process of claim 1 in which the brine leach
65 comprises an aqueous solution containing about
200-300 gpl to saturation of a soluble chloride which
provides a maximum concentration of chloride ion
below saturation.
4,556,422
9
5. The process of claim 4 in which the chloride is a
member selected from the group consisting of chlorides
of alkali and alkaline earth metals and hydrogen chloride.
6. The process of claim 1 in which the chloride solu- 5
bilized is lead chloride.
7. A process for selectively solubilizing a chloride
selected from the group consisting of lead cWoride and
a combination of lead and silver chlorides contained in
a mixture of other solids including metal sulfides which 10
comprises subjecting the mixture to a brine leach under
a normal boiling point.
8. The process of claim 7 in which the temperature is
in excess of about 100· C.
9. The process of claim 8 in which the leaching is 15
performed in a time not in excess of about three minutes
and a concentration of at least about 130 gpl of lead
chloride is obtained.
10. The process of claim 8 in which said metal sulfides 20
include the sulfides of copper, iron and zinc.
11. The process of claim 10 in which the time of leach
is less than about three minutes and a concentration
tration of at least about 130 gpl of lead chloride is obtained.
25
12. The process of claim 10 in which the brine leach
comprises an aqueous solution containing at least about
200 gpl of soluble chloride which provides a maximum
concentration of chloride ion below saturation.
13. The process of claim 12 in which said cWoride is 30
a member selected from the group consisting of chlorides
of alkali and alkaline earth metal cWorides and
hydrogen cWoride.
14. The process of claim 8 in which the chloride
solubilized is lead cWoride. 35
15. The process of claim 13 in which said chloride is
sodium cWoride.
16. The process of claim 8 in which the solubilized
silver and lead cWorides are separated from solids.
17. The process of claim 16 in which said separation 40
is accomplished by liquid cyclone separation.
10
18. :The process of claim 16 in which lead chloride is
crystallized from the solution and silver chloride is
recovered from the mother liquor.
19. The process of claim 16 in which lead chloride is
recovered from the brine solution by crystallization.
20. The process of claim 19 in which said crystallization
includes flashing from the brine leach temperature
to a lower temperature.
21. The process of claim 8 in which said mixture
includes elemental sulfur and the temperature of said
brine leach is at the agglomeration temperature of sulfur.
22. The process of claim 19 in which lead is recovered
from the crystallized lead cWoride.
23. The process of claim 22 in which the lead is recovered
by hydrogen reduction of the crystallized lead
chloride.
24. The process of claim 23 in which the hydrogen
and the heat requirement for the endothermic reduction
of lead chloride are supplied by a partially combusted
hydrocarbon gas.
25. The process of claim 23 including supplying the
heat required for the endothermic reduction by heating
molten lead introduced into the reaction area.
26. A process for solubilizing a cWoride selected from
the group consisting of lead chloride and a combination
of silver and lead chlorides which comprises subjecting
the chloride to a brine leach at a temperature above the
normal boiling point under a pressure above atmospheric
sufficient to prevent boiling for a time not in
excess of 3 minutes sufficient to solubilize substantially
all of the chloride.
27. A process for solubilizing a cWoride selected from
the group consisting of lead cWoride and a combination
of silver and lead chlorides which comprise subjecting
the cWoride to a brine leach at a temperature above the
normal boiling point under a pressure above atmospheric
sufficient to prevent boiling for a sufficient time
to obtain a concentration of at least about 130 gplof
solubilized lead cWoride. • • • • •
45
50
55
60
65
.
0.022
0.015
0.020
0.D15
0.D18
0.017
0.023 .
0.018
0.012
0.013
0.010
0.112
0.026
0.019
0.013
0.012
0.010
0.009
5
10
15
30
60
90
5
10
15
30
60
o5
10
15
30
60
90
25
~
~
~
~
I~
In the third stage of the process of the present invention,
pulp from the pressure oxidation step is passed
through a surge tank to which a basic chemical may be
35 added for pH adjustment. It has been found that the
pulp from the pressure oxidation step should preferably
be cooled to a temperature below about 50· C. before it
is introduced to carbon-in-Ieach treatment. This cooling
may take place in a heat exchanger which recovers part
of the heat to be used upstream from the heat exchanger.
In addition to cooling the pulp, it may be
necessary to dilute the pulp before the carbon-in-Ieach
treatment wherein cyanide may be added to the first of
a plurality of mechanically agitated vessels in series in
which gold extraction from ore by cyanidation and
carbon absorption will proceed simultaneously. Pulp is
transferred continuously downstream through interstage
screens from a first vessel to the following vessel
in the series of vessels while activated granular charcoal
carbon is advanced from the last vessel toward the first
vessel. Fresh reactivated granular charcoal carbon is
added to the last stage and the loaded charcoal is withdrawn
from the first stage. Pulp leaving the last stage is
passed through an additional screen to scavenge some
attrited carbon before being discarded to a tailings pond
from which the tailings water may be recycled to the
slurry formation step.
When liquid from the tailings pond is recycled upstream
of the autoclaves, it has been found that free
cyanide in the recycled liquid unexpectedly decreases
the efficiency of the gold recovery of the present invention.
This is in contrast to conventional cyanidation
techniques in which the excess free cyanide would in
fact be utilized to help oxidize the slurry to be treated.
To overcome this unexpected problem, the free cyanide
must be removed by any suitable means.
The invention will be further illustrated in the following
example in which the pressure oxidation step is
20 Composite DC-I:
Composite DC-2:
25
Temp 0c.
180 50 0.028
200 ~ 0.020
215 ~ 0.017
225 ~ 0.016
249 ~ 0.021
250 ~ 0.016
150 100 0.070
180 ~ 0.031
180 ~ 0.027
215 ~ 0.021
225 ~ 0.011
Composite 2:
150 25 0.020
180 ~ 0.017
225 ~ 0.007
Composite DC-2:
150 0.060
180 0.023
200 0.014
215 0.014
225 0.010
All tests at 60 minutes and 40% solids.
In order to achieve adequate mixing without unnecessarily
diluting the slurry, it has been found preferable
to maintain the solids content of the slurry between
about 40% to about 50% and it is postulated that the
rate of reaction in the pressure oxidation stage is con- 40
trolled by the mass transport of oxygen to the solids'
surface. However, the selection of the proper type of
mixing equipment and the mixing speed utilized within
the autoclaves is deemed to be well within the scope of
one of ordinary skill in the art. 45
Another critical variable in the pressure oxidation
step is the residence time which the ore slurry spends in
the autoclaves. By way of example only, a summary of
tests evaluating the effect of autoclave time is given in
Table 7 and shown graphically in FIG. 2. The data 50
show a curve with minimum slope from approximately
30-90 minutes. Progressively higher tails assays result
from shorter times. Complete oxidation of the sulfides is
unnecessary; within an average 30-minute treatment
period found to be adequate, less than 50% of the sul- 55
fides may be oxidized. To maintain the oxygen pressure,
a bleed is required from the autoclaves as carbon dioxide
is evolved from the reaction of carbonates noted
above. Additionally, since the amount of oxygen consumed
will inevitably vary depending upon the refrac- 60
tory nature of the ore being treated, the oxygen partial
pressure must be monitored to assure it does not drop
below a minimum of about 10-25 psia. The partial pressure
of oxygen may be maintained by the introduction
of either pure oxygen, air or a mixture of both into the 65
autoclaves to ensure that an effective amount of the
substances which cause the ore to be refractory will be
oxidized in the pressure oxidation step.
4,552,589
60
11
described in greater detail. Thus, the operating conditions
in the autoclaves are as follows: a temperature of
about 2250 C., a retention time of about 30 minutes, an
oxygen overpressure· of about 25 psia, an agitation of
about 100 hp/5,OOO gallon unit, a pulp density of about 5
45% solids, a grind of about 80% passing through 325
mesh, a reagent addition of about 8 pounds NaOH/ton
ore, an oxygen supply (average) of about 45 lb/ton ore
and a steam supply (maximum) of about 250 1b/ton of
ore. At least six stages are utilized to avoid by-passing 10
inefficiency. This might be achieved through six vertical
units or one or two multistage horizontal autoclaves.
The gas bleed and oxygen input are controlled by manifolding
the bleed gas from each stage and exhausting to
control pressure, while injecting oxygen to maintain a 15
desired oxygen concentration in the atmosphere of each
vessel. The agitation is designed to reentrain gas at the
pulp surface. A flash heat exchange system is used to
recover heat, followed by steam injection to attain the
desired pulp temperature in the first stage. Bleed gas is 20
contacted with feed pulp to recover some heat as well
as to provide initial scrubbing. Although the pulp
would be alkaline, pitting corrosion could occur above
the pulp interface. Accordingly, a mild steel autoclave 25
would require at least a partial protective lining.
Having fully described the present invention, it will
be apparent from the above description and drawings
that various modifications in the process of the present
invention may be made within the scope of this inven- 30
tion. Therefore, this invention is not intended to be
limited except as may be required by the lawful scope of
the following claims.
What is claimed is:
1. A process for recovering gold from a refractory 35
ore slurry, said process comprising the steps of:
introducing the slurry to at least one agitated autoclave
maintained at an autoclave temperature in
excess of about 1500 C. and an autoclave oxygen
partial pressure in excess of about 10 psia; 40
agitating the slurry to allow oxygen to oxidize at least
part of the substances which cause the ore to be
refractory to form an oxidized slurry;
passing the oxidized slurry to a plurality of stages in
which the oxidized slurry is contacted with a cya- 45
nide and activated charcoal, the charcoal flowing
in countercurrent fashion with said oxidized slurry
whereby precious metals are transferred to the
activated charcoal; and
separating the activated charcoal from the oxidized 50
slurry.
2. The process as recited in claim 1 wherein the autoclave
temperature is maintained between about 1800 C.
to about 2250 C.
3. The process is recited in claims 1 or 2 wherein the 55
autoclave oxygen partial pressure is maintained at a
minimum between about 10 psia and about 25 psia.
4. A process as recited in claim 1 wherein a basic
compound, is added to the ore slurry to form an insoluble
species with free sulfate ions.
5. A process as recited in claim 4 wherein the basic
compound comprises sodium carbonate.
6. A process as recited in claim 1, comprising the
further step of:
cooling the oxidized slurry before the oxidized slurry 65
is contacted with the cyanide.
7. A process as recited in claim 1, comprising the
further step of:
12
raising the pH of the oxidized slurry before the oxidized
slurry is contacted with the cyanide.
8. A process as recited in claim 1, comprising the
further step of:
diluting the oxidized slurry before the oxidized slurry
is contacted with the cyanide.
9. A process as recited in claim 1 wherein the ore
slurry is maintained within the autoclave step from
approximately 30-90 minutes.
10. A process for recovering gold, said process comprising
the steps of:
forming an ore slurry from a refractory gold bearing
ore;
introducing the slurry to to at least one autoclave
which is maintained at an autoclave temperature in
excess of about 1500 C. and an autoclave oxygen
partial pressure maintained in excess of about 10
psia;
agitating the slurry sufficiently to oxidize at least part
of the substances which cause the ore to be refractory
to form an oxidized slurry;
passing the oxidized slurry to a plurality of stages in
which the oxidized slurry is contacted with a cyanide
and activated charcoal at a cyanidation temperature
less than about 500 C., said charcoal flowing
in countercurrent fashion with said oxidized
slurry whereby gold is transferred to the activated
charcoal; and
separating the activated charcoal from the oxidized
slurry.
11. The process as recited in claim 10 wherein the
autoclave temperature is maintained between about
1800 C. to about 2250 C.
12. The process as recited in claims 10 or 11 wherein
the autoclave oxygen partial pressure is maintained at
about 15 psia.
13. A process as recited in claim 10 wherein a basic
compound is added to the ore slurry to form an insoluble
species with free sulfate ions.
14. A process as recited in claim 13 wherein the basic
compound is selected from the group consisting of sodium
carbonate, sodium bicarbonate and sodium hydroxide.
15. A process as recited in claim 10 wherein the gold
in the ore exists in a finely disseminated form with an
average size of less than 10 microns.
16. A process as recited in claim 10 wherein the refractory
ore contains a sulfur bearing compound which
is at least partially oxidized in at least one autoclave.
17. A process as recited in claim 10 wherein the ore is
less than 60% amenable to standard cyanidation techniques.
18. A process as recited in claim 17 wherein more
than 85% of the gold existing in the ore slurry is transferred
to the activated charcoal.
19. A process as recited in claim 10 wherein the ore
slurry is maintained within at least autoclave for from
about ten minutes to about thirty minutes.
20. A process as recited in claim 10 wherein the activated
charcoal is separated from the oxidized slurry
from about 4 hours to about 8 hours after said activated
charcoal is contacted with the oxidized slurry.
21. A process for recovering gold, said process comprising
the steps of:
forming an ore slurry by grinding a refractory gold
bearing ore to form a slurry which is mixed with a
diluent stream;
4,552,589
13
introducing the slurry to at least one autoclave which
is maintained at an autoclave temperature in excess
of about 150· C. and an autoclave oxygen partial
pressure maintained in excess of about 10 psia;
agitating the slurry to oxidize substances which cause 5
the ore to be refractory to form an oxidized slurry;
cooling the oxidized slurry to a temperature less than
about 150· C.;
passing the oxidized slurry to a plurality of stages in
which the oxidized slurry is contacted with a cya- 10
nide and activated charcoal, the activated charcoal
flowing in countercurrent fashion with the oxidized
slurry whereby gold is transferred to the
activated charcoal; 15
separating the activated charcoal from the oxidized
slurry;
collecting a tailings stream from the oxidized slurry;
treating a water portion of the tailings stream to destroy
any free cyanide and form a treated reclaimed 20
water stream; and
recycling the treated reclaimed water stream to a step
ofthe process upstream from the step of passing the
oxidized slurry to a plurality of stages.
14
22. A process as recited in claim 21 wherein the
treated reclaimed water stream is combined with a feed
stream to form the diluent stream.
23. A process as recited in claim 21 wherein the oxidized
slurry is cooled by a heat exchanger which is
utilized to preheat the slurry.
24. A process as recited in claim 21, comprising the
further step of:
diluting the oxidized slurry to dethicken said oxidized
slurry before the oxidized slurry is contacted with
the cyanide.
25. A process as recited in claim 24 wherein the reclaimed
water stream is used to dilute the oxidized
slurry.
26. A process as recited in claim 21 wherein the ore
slurry is maintained within at least one autoclave from
about 10 minutes to less than about 60 minutes and the
carbon-in-leach residence time is from about 4 hours to
about 8 hours.
27. A process as recited in claim 26 wherein the ore
slurry is less than 60% amenable to standard cyanidation
techniques and more than 85% of the gold in said
ore slurry is transferred to the activated charcoal.
* * * * *
25
30
35
40
45
50
55
60
65
ginati} ��oe@s� ut-grid-align:none;text-autospace:none'>portion of said high pressure leachate of step (c) and at
least a portion of said second fraction of step (a), are
contacted in an atmospheric leach at atmospheric pressure
and at a temperature below about 100· C. to form 20
an atmospheric leach residue and an atmospheric leachate
prior to contact in the low pressure leach of step (d).
4. A method according to claim 3 further comprising
separating from said first fraction a coarser third frac- 25
tion, said third fraction containing more magnesium
than said first fraction and less than said second fraction,
and contacting at least a portion of said third fraction
and at least a portion of said atmospheric leachate in the
low pressure leach of step (d). 30
5. A method according to claim 1 further comprising
separating from said first fraction a coarser third fraction,
said third fraction containing more magnesium
than said first fraction and less than said second fraction,
and contacting at least a portion of said third fraction in 35
the low pressure leach of step (d).
6. A method according to claim 1 or claim 3, wherein
at least a portion ofsaid low pressure leachate is neutralized
by the addition of a neutralization agent selected
from the group consisting of alkali and alkaline earth 40
oxides and alkali and alkaline earth hydroxides to form
a neutralized low pressure leachate.
7. A method according to claim 1, or claim 2, or claim
3, or claim 4, wherein at least a portion of said low
pressure leach residue is recycled to said high pressure 45
leach.
8. A method according to claim 3 or claim 4, wherein
at least a portion of said low pressure leach residue is
recycled to said atmospheric leach.
9. A method according to claim 8 wherein a coarser 50
fraction of said atmospheric leach residue is separated
from said atmospheric leachate and the remainder of
said atmospheric leach residue, and further comprising
filtering said coarser fraction into a filtrate and a filter
cake from which chromite is recovered.
10. A method according to claim 9 wherein at least a
portion of said filtrate is recycled to said atmospheric
leach.
60
65