United States Patent Office 3,477,928
Patented Nov. 11, 1969
1 2
It is still another object of the invention to provide
an economic process for recovering improved yields of
silver from ores containing such small amounts that
other well known processes have been heretofore consid·
5 ered impractical.
These and other related objects are achieved by a process
which comprises a combination of steps or operations
which provide conditions and accomplish results essential
to the success of the overall process.
10 Broadly, the instant process comprises mixing a comminuted
ore material with an effective amount of a
chloridizing agent; roasting the' resulting admixture under,
first, non-oxidizing conditions,. and then, under oxidizing
conditions; leaching the resulting calcine in a brine solu-
15 tion to extract the metal values from the calcine; and then
recovering the metals by electrodeposition techniques.
Oxidized silver-lead ores to which the invention is particularly
adapted include low-grade ores consisting of siliceous,
ferruginous gossan containing at least about 1
20 weight percent lead, from about 6 to about 30 weight
percent iron, from about 0.5 to about .5 weight percent
sulfur, and about 3 to about 20 ounces of silver per ton
of ore. The analysis of illustrative ores are set forth
below:
The raw ore is first ground to a suitable particle size
such that it can be intimately and uniformly mixed with
the chloridizing agent. Generally, the are should be
ground to such a size that from about 50 to about 100%
of the ground ore' will pass through a -325 mesh screen.
30 Preferably, at least about 70% of the ore should pass
through a -325 mesh screen. Experiments have indicated
that the extraction of lead is not critically dependent
upon the fineness of the are particle'S. The extraction of
silver, however, is significantly affected! by the particle
ABSTRACT OF THE DISCLOSURE
This invention relates to a process for the recove'ry
of metals. More specifically, the invention relates to a
process for the recovery of silver, lead, copper and bis-
3,477,928
PROCESS FOR THE RECOVERY OF METALS
Enzo L. Coltrinari, Arvada, Colo., assignor to Cerro
Corporation, New York, N.Y., a corporation of New
York
No Drawing. Filed Mar. 28, 1966, Ser. No. 537,722
Int.CI. C22d 1/26
U.S. CI. 204-123 7 Claims
An improved, continuous process for the recovery of
lead, silver, copper and bismuth metal values from lowgrade,
oxidized lead ores containing said metal values,
which comprises comminuting at least about 50 to 70
percent of the ore material to minus 325 mesh size, mixing
the comminuted are with an effective amount of a
chloridizing agent sufficient to convert the metal values
into brine-soluble' form, roasting the resulting admixture
in a two-stage, nonvolatilizing roast operation, the first
stage being conducted at a temperature between about
5100 C. to about 6600 C. under a neutral atmosphere
to remove a substantial portion of any sulfur originally
present in the are, the second stage being conducted at
a temperature between about 5000 C. and about 540 0 C.
under an oxidizing atmosphere to convert the metal values 25
into brine-soluble form, leaching the resulting calcine
in a brine solution with a pH between about 1.9 and about
2.5 to extract the metal values therefrom and subsequently
recovering the metals by electrodeposition techniques.
TABLE 1
Percent
Ore Oz.jton Cn Pb Zn Bi 8iO' Fe S 804/8 H,O
A_________ B_________ 7.57 0.23 1. 08 0.3 0.11 61. 5 14.2 3.86 0.51 5.64
C_________ 10.50 0.10 1.10 O. b 0.03 72.5 9.5 1.82 0.73 5.91 D _________ 8.35 0.09 4.00 0.5 0.03 60.6 15.5 1. 76 0.53 6.06 E _________ 6.30 0.13 2.10 0.5 0.06 67.5 11.8 1.37 0.50 3.44 8. 80 0.38 3.90 2.0 0.10 33.0 26.9 0.70 0.16 16.75
TABLE 2
size of the ore. The influence of particle size' on the re55
covery of silver and lead is shown in Table 2 below:
8ilver Lead
Total extraction (percent)
88.5
89
88
88
87
90
94
95;
Percent of ore, -325 mesh
60 45 •• _•• _
8~8£=====:•======::=====::====:=•===_
muth from low grade oxidized lead ores containing said
metals.
It is an object of this invention to provide an efficient
process for the recovery of silver, from low-grade silverlead
ores in which the silver is present in comparatively
minor quantities.
It is another object of the invention to provide a process
characterized by improved recovery of silver and lead
from low-grade silver-lead ores which process can be conducted
in a batch-wise or continuous operation.
3,477,928
4
present during the roast. The temperature during the first
stage roast is preferably maintained between about 510
and 6600 C. and more preferably at about 560 0 C. If
the temperature of this first stage roast is increased more
5 of the sulfur can be removed and the extraction of lead
can be markedly increased. However, high temperatures
have been found to be detrimental to the recovery of
silver. Accordingly, a practical balance between silver
recovery and sulfur removal must be accepted. Any sulfur
10 which remains in the ore will be converted to sulfate during
the second stage roast and will be extracted into the
brine during the leaching step. This sulfate can be controlled
by adequate treatment of the brine as de'scribed
herein.
The second stage roast, which is conducted under oxidizing
conditions, is directed to the conversion of the lead,
silver, copper and bismuth to a form which can be easily
and efficiently leached from the roasted ore by a brine
solution. A suitable oxidizing atmosphere can be pro-
20 vided by the use of air, alone or in admixture with the
normally present combustion gases. It has also been
found beneficial to employ small amounts of chlorine
gas, e.g., up to about 5 percent by volume of the atmosphere,
during the second stage of the roasting operation.
25 Chlorine gas has been found to be particularly beneficial
in connection with the lead extraction.
The second stage chloridizing roast under an oxidizing
atmosphere is conducted at temperatures which are sufficient
to convert the metal values to a brine leachable
30 form without volatilization. Generally, this chloridizing
roast is conducted at temperatures from about 500 0 C. to
about 5400 C. and preferably about 510 0 C. The time
required to remove the sulfur and to convert the metal
values to a brine soluble form can be easily determined
35 by analysis of the gas removed during the first stage roast
and analysis of the brine leach liquor for silver and lead.
It has been found that about a 30-minute first-stage roast
is sufficient for complete removal of the sulfur and
about a 30-minute second-stage roast is sufficient for re-
40 covery of about 92% of the silver. Maximum lead recovery
can be obtained after an oxidizing roast of about
15 minutes.
After roasting the resulting calcine is mixed with a
concentrated acidified brine solution to form a slurry
45 in order to extract the chloridized metals from the calcined
ore. The amount of brine used per ton of raw ore
depends at least in part upon the lead content of the
original ore. Generally, the solids content of the ore slurry
should be from about 25 to about 40 percent and pref-
50 erably about 33 percent. The amount of brine actually
employed should be an amount sufficient to provide a
leach liquor having a lead content of from about 15 to
about 20 grams per liter. Accordingly with high content
lead ores a slurry of lower solids content should be
55 employed; conversely if a low grade ore is used, the
proper lead concentration can be provided by using less
brine per tone of ore thereby providing a slurry of relatively
high solids.
Solubility of the metal values in the brine is effected
60 by the salt concentration in the brine as well as the temperature
and acidity thereof. Moreover, because lead sulfate
has a low solubility, the concentration of sulfate ion
in the brine must be limited, as mentioned above, to prevent
loss of lead by precipitation during leaching.
The solubility of lead and silver increases significantly
with increase in the sodium chloride concentration of the
brine leach liquor until the saturation point of sodium
chloride is reached. At this point, the solubility drops
abruptly. Accordingly, a concentrated but non-saturated,
70 brine having a sodium chloride content of between about
200 and 300 grams per liter of sodium chloride and preferably,
from about 272 to about 297 grams per liter of
sodium chloride is generally employed. Since additional
chloride from the roasted ore is introduced into the brine,
75 it may be necessary in recycle operations to dilute the
Ag
52.6
68.6
91. 8
94.3
Pb
79.9
82.6
83.5
83.1
Extraction (percent)
NaC! (percent)
3
After grinding, the ore is intimately and homogeneously
mixed with an effective amount of a chloridizing agent,
preferably sodium chloride, to provide a homogeneous
admixture suitable for roasting. The term "effective
amount" is intended to refer to an amount of chloridizing
agent which is sufficient to convert the metal values in
the ore to a brine-soluble form. Generally, from about
3 to about 7 and preferably from about 3.5 to 5 weight
percent of sodium chloride is employed as the chloridizing
agent. Variation in the amount of sodium chloride has
been found to have a more significant effect on the recovery
of silver than upon the recovery of lead. The
sodium chloride can be added to the ore in dry form
either before or after grinding, or it can be added to the
ore in the form of a concentrated aqueous solution. It is 15
believed that the sodium chloride has an important effect
on the solubilization of copper, zinc and bismuth. Generally,
the sodium chloride should be ground to a particle
size of at le'ast about -65 mesh since the particle
size of the chloridizing agent has been found to have a
substantial effect on the recovery of silver.
The effect of varying concentrations of sodium chloride
on recovery of silver and lead is shown in Table' 3,
below. In these investigations, the ore was mixed with
the indicated amount of -65 mesh sodium chloride
and then roasted in air at 5100 C. for a period of about
sixty minutes. The calcined ore was then leached with
acidified brine containing 270 grams per liter of sodium
chloride. Analysis of residue and of the leach liquor gave
the results tabulated in Table 3.
TABLE 3
0 _
L _
3 _
5 _
In addition to serving as the chloridizing agent, the
sodium chloride also influences the removal of sulfUr
from the ground ore during the first stage, Le., the nonoxidizing
stage, of the roasting operation. Investigation
has shown that the ratio of sulfate to lead in the brine
leach liquor can be controlled, at least in part, through
the amount of sodium chloride employed, and is significantly
decreased by increasing the amount of sodium
chloride from about 2 to about 7% by weight based on
the weight of the ore.
The mixture of ore and sodium chloride is roasted at a
temperature which is sufficient to convert the metal values
to a brine soluble form without undue loss of metal
values through volatilization. Roasting of the ore and
sodium chloride is conducted in a two-stage operation.
The first stage, intended primarily for the removal of a
substantial portion of the sulfur originally contained in
the ore, is preferably conducted under a neutral atmosphere,
e.g., in an atmosphere comprising nitrogen, carbon
dioxide and water vapor preferably in a volume ratio
of about 8: 1: 1. The term "neutral" used herein refers
to an atmosphere which favors the conversion of sulfur
to a form which can be removed by volatilization, I.e.,
an atmosphere which will neither oxidize sulfur to sulfate
nor reduce volatile sulfur compounds to a non-volatilizable
form. Removal of sulfur is important since the presence
of oxidizable sulfur in the ore during the second 65
stage roast results in an increase in the concentration of
sulfates ultimately appearing in the brine liquor after
leaching.
In the two-stage roasting operation, the sulfur removal
stage is carried out at a temperature which is sufficient to
remove substantially all of the sulfur originally present
in the ore, but without volatilization of any of the metal
values. The amount of sulfur removed is dependent upon
the temperature of the first stage roast as well as upon
the atmosphere and the amount of chloridizing agent
3,477,928
Ag (Percent) Pb (Percent) 40
H CI/ton of ore ~% NaC! 3.~% NaCl ~%NaCl 3.5% NaCl
(Percent) Roast Roast Roast Roast
10 90 87. ~ 84 81
20 92 90 86 84
40 93 91 89.5 88 45
6
oxidizing roast would result in a reduction of the overall
cost of operation.
After leaching, the brine leach liquor is separated from
the ore by filtration or some other suitable method. In
5 order to properly separate all the dissolved metal values
from' the leached ore the ore should be washed, e.g., by
repulping and refiltering, in order to dilute all of the dissolved
metal values and transfer them to a suitable
aqueous solution for subsequent recovery of the metal
10 values by electrolysis.
The metal values, including the silver, copper, and bismuth,
are recovered from the leach liquor, in the form
of a' lead bullion by electro·deposition techniques. The
lead bullion thus provided can be further processed for
15 the recovery of the individual metals by various metal·
lurgical techniques well known in the art, e.g., fire or
electrolytic refining. Generally, the lead bullion is deposited
on a suitable cathode, e.g., a cathode of lead or
unpolished iron, in the form of relatively dense adherent
20 sponge by electrolyzing the leach liquor under suitable
conditions of temperature and current density, preferably
at a temperature at least about 70° C., a ,current density
between 'about 12.0 and 19.5 amperes per square foot,
and a -voltage between about 2.0 and 2.4 volts.
Employment of temperatures of at least about 70· C.
provides a suitable lead deposit which is easy to handle
and which adheres to the cathode without falling or flaking
off. Moreover, such temperatures provide for more
efficient power consumption and depletion of the lead
30 content of the brine to a low level, i.e. about 0.2 gram per
liter.
During electrolysis, the brine solution should be agitated
with sufficient vigor to circulate the brine without
causing the deposited lead to fall from the cathode. Agi-
35 tation was found to significantly effect the characteristics
of the lead bullion deposited on the cathode. If agitation
was too gentle, the cell resistance was found to increase,
resulting in a higher voltage and increased gas formation.
The increased amount of gas formation was found to
cause a more spongy and less adherent deposit.
During the electrolytic precipitation of lead, silver,
copper and bismuth from the brine leach liquor as described
above, a brown precipitate forms in the electrolytic
cell. This material is comprised mainly of manganese
dioxide as well as small amounts of compounds
of other metals such as aluminum, iron, chromium, silicon,
magnesium, calcium, sodium and the like. This material
may be separated by filtration of the leach liquor.
Similarly, zinc· may be recovered by the precipitation of
zinc hydrate from the brine after electro·deposition by
treating brine wIth sodium hydroxide or lime, and separating
the precipitate by filtration, or other equivalent
means.
The process described herein, can be carried out in a
55 batch-wise or a continuous process in which the depleted
leach liquor after electrolysis may be recycled after suitable
treatment to remove precipitated manganese dioxide
and for sulfate control and used for the subsequent
leaching of additional batches of calcined ore.
I claim:
1. A process for recovering lead, silver, copper and
bismuth metal values from a low-grade, oxidized lead
ore containing said metal values which comprises the
following steps:
(1) Comminuting the original ore to such a particle
size that from at least about 50 to about 70 percent
by weight of said ore will pass through a minus 325
mesh screen;
(2) mixing said comminuted ore with an effective
amount of sodium chloride, ground to a particle size
of at least about minus 65 mesh, sufficient to convert
the said metal values into brine·soluble form;
(3) roasting the resulting admixture in a two-stage,
nonvolatilizing roast operation, the first stage being
conducted at a temperature ranging from about 510·
10
130
45
92.0
85.8
2
40 40
68 130-150
60 40
92.7 92.9
90.1 90.8
Extraction
As mentioned above the level of sulfate ion concentration
in the brine roust be controlled. If the concentration
of sulfate in the brine reaches high enough levels
lead will precipitate in the form of lead sulfate thereby 60
reducing the recovery of lead. Increasing concentrations
of sulfate ions is a particularly serious problem when a
continuous process comprising recirculation of the brine
is employed. It will be appreciated that the presence of
excessive amounts of sulfate in the brine will increase the 65
cost of controlling the level of sulfate in the recirculating
brine. Generally, the sulfate concentration is controlled
by precipitation of the excess sulfate as calcium sulfate
thrpugh the addition of calcium chloride to the brine.
The calcium chloride can be added directly to the brine 70
orit can be provided, in situ, by adding hydrochloric acid
and lime to the brine. Alternatively, the sulfate concentration
can be controlled by refrigeration of the brine,
thereby precipitating the sulfate as hydrated sodium suI·
fate. Obviously, the removal of any sulfur prior to the 75
HCI Added, Lbs. 100% RC./ton calcine __•
Leach temperature,· C•• •• ••
Leach tims, Mins._••••• ._ •• __ ••_•• __ ••
Extraction:
Percent Ag••_•• • • ••__ ••
Percent Pb • •
Table 5 below shows the influence of temperature upon
the extraction of silver and lead.
TABLE 5
5
brine in order to maintain these optimum levels of sodium
chloride.
Hydrochloric acid .is added to the brine leach liquor
during the leaching operation. This acid is required partly
to neutralize the lime which is added for the purpose of
sulfate control and partly to improve extraction of the
metal values from the ore. It has been found experimentally
that increasing the acidity of the leach liquor increases
the extraction of silver and lead.
With respect to silver and lead, it has been observed
that variations in the quantity of sodium chloride added
to the roast is related to the acidity necessary for efficient
leaching. For example, if 40 pounds of hydrogen chloride
are used for leaching a calcined ore, increasing the
amount of sodium chloride addition from 3.5% to about
7% by weight does not influence the extraction of lead.
However, if smaller amounts of hydrogen chloride, e.g.,
about 10 pounds, is used to acidify the brine, the amount
of sodium chloride employed has a considerable influence
on subsequent extraction of lead. In the case of silver, reduction
of the amount of sodium chloride used during
roasting of the ore, results in a decrease in the silver extraction,
even if 40 pounds of hydrogen chloride per ton
of ore are added to the brine for leaching. In general, the
brine leach liquor should be acidified with between 15 25
and 40 pounds of hydrogen chloride per ton of ore in
addition to the amount of hydrogen chloride which is required
to provide calcium chloride for sulfate control.
The pH of the brine solution should be between about
1.9 and 2.5 and preferably should be approximately 2.2.
A pH of about 2.2 can be provided by adding about 20
pounds of hydrogen chloride per ton of ore to the brine
liquor in addition to the 'amount required for maintaining
the sulfate at an operable concentration level. The influence
of variations in acidity and in the amount of salt
employed during the roast, is shown in Table 4 below.
TABLE 4
3,477,928
References Cited
UNITED STATES PATENTS
9/1907 Davis 75-7
8/1909 Eldred 75-7
11/1918 Larson 204-117
2/1921 Bradford 75-113
11/1934 Levy 204-117 XR
866,580
932,689
1,284,910
1,368,885
1,980,809
U.S. Cl. X.R.
75-7;204-105,106,107,109,111,114,117
25 JOHN H. MACK, Primary Examiner
G. L. KAPLAN, Assistant Examiner
8
deposition of said lead bullion is carried out at a temperature
of at least 700 C., a current density of about 12.0
to about 19.5 amperes per square foot and a voltage of
about 2.0 to about 2.4 volts.
5 6. The process according to eIaim 1 wherein an effective
amount of sodium chloride is from about 3 to about
7 percent by weight of the comminuted ore.
7. The process according to claim 1 wherein said metal
values are deposited as lead bullion in a relatively dense,
10 adherent sponge like form on a cathode consisting essentially
of a metal selected from the group consisting of
lead and unpolished iron, said brine solution being agitated
during electrolysis with sufficient vigor to circulate
the brine without causing the deposited lead bullion to
15 fall from said cathode.
7
c. to about 660 0 C. under a neutral atmosphere to
remove a substantial portion of any sulfur originally
present in said ore, and the second stage being conducted
at a temperature ranging from about 5000 C.
to about 5400 C. under an oxidizing atmosphere to
convert said metal values to brine-soluble form;
(4) leaching the resulting calcine in a brine solution
with a pH between about 1.9 to about 2.5 to facilitate
extraction of said metal values from said calcine;
and
(5) subsequently recovering said metal values by e1ectrodeposition
techniques.
2. The process according to claim 1 wherein the neutral
roast stage of the roasting operation is conducted at
about 5600 C. in an atmosphere comprising nitrogen,
carbon dioxide and water vapor in a volume ratio of about
8: 1: 1 and wherein the oxidizing stage of said roasting
operation is conducted at about 5100 C. in the presence
of air.
3. The process according to claim 2 wherein the at- 20
mosphere under which the second stage of the roast is
conducted comprises air and up to about 5 percent by
volume of chlorine gas.
4. The process according to claim 1 wherein the calcined
ore is leached with a concentrated, but unsaturated
brine solution containing from 200 to 300 grams per liter
of sodium chloride to form a slurry containing from
about 25 to about 40 percent solids.
5. The process according to claim 4 wherein electro