United States Patent [19)
Holmes et ale
[ II]
[45)
3,709,680
Jan. 9, 1973
FOREIGN PATENTS OR APPLICATIONS
Primary Examiner-Allen B. Curtis
Attorney-March and Le Fever
1899 Great Britain 75/118
[54) PROCESS FOR REMOVAL OF
ARSENIC FROM SULFO·ORE
[75] Inventors: W. Church Holmes, Lake San Marcos,
Calif.; Enzo L. Coltrlnari, Arvada,
Colo.
[73] Assignee: Sunshine Mining Company, Kellog,
Idaho
[22] Filed: July 9,1971
[21] AppI. No.: 161,165
2,835,569
3,218,161
3,476,553
22,619
[57]
5/1958
11/1965
11/1969
Reynaud et al... 75/118
Kunda et ai. 75/118 X
Sebbaetai. 75/121 X
ABSTRACT
[52] U.S. CI 75/6, 75/101 R, 75/118,
75/121
[51 ] Int. CI. C21b 1/04, C22b 61/00
[58] Field ofSearch 75/101 R, 118, 121,6
[56] References Cited
UNITED STATES PATENTS
683,325 9/1901 PhiIlips , 75/118
726,294 4/1903 Hoyt... 75/118
A process and apparatus for the removal of arsenic
values from an ore concentrate includes the steps of
leaching the concentrate with a solution which dissolves
the arsenic, separating undissolved residue
therefrom, acidifying the arsenic pregnant solution to
precipitate arsenic and other mineral values as insoluble
salts and treating the arsenic barren solution to
regenerate sulfur and sodium values for recycle to the
process.
10 Claims, 2 Drawing Figures
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HZs----+----
S02 TO RECYCLE
NA'... S Ttl R,tCYCLE
PATENTEOJAH' 91973 3.709.680
SHEET 1 OF 2
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PATENTEOJAH 9 1973
SHEET 2 OF 2
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INVENTORS
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BY
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3,709,680
2
35
The temperature of the concentrated sodium sulfide
solution used to dissolve the arsenic from the flotation
concentrate must be carefully controlled in order to
obtain the optimum results. The sodium sulfide solu-
5 tion normally contains from about 100 to about 600
grams per liter of sodium sulfide, sodium sulfide-sodium
hydroxide or the equivalent of the other mixtures.
The temperature of the leaching solution in order to
obtain the maximum results in an efficacious and rapid
10 manner should be within the range at sea level of from
about 75°C to the temperature just below boiling,
preferably from about 100° to about 205cC. It has been
found that optimum extraction of arsenic from the
flotation concentrate occurs if the leaching time is
15 maintained at one within a range of from about I to 20
h~urs, preferably from about 4 to 12 hours. Agitation is
preferably maintained to aid in insuring an adequate
contact between the leaching solution and the flotation
20 concentrate.
The major content of the concentrate remains in the
residue as the corresponding sulfides, and is separated
in that form for value recovery.
The hot leach mixture from the leaching operation is
25 diluted with preferably hot water prior to separation of
residues. The dilution prevents undesirable crystallization
of the salts with temperature drop and results in
more efficient and effective further processing.
The diluted mixture is preferably dewatered by a plu-
30 rality of thickeners, arranged for counter-current decantation,
the thickened slurry filtered and the filtrant
contains desirable mineral values. The filtrate is combined
with no decant from the thickening step and
passed to the next stage of the process.
As an example of the leaching step the following is
given.
A sample of a flotation concentrate containing 12.2
percent arsenic and weighing 100 grams was contacted
with a solution containing 390 grams of sodium sulfide
per liter of water. The solution was agitated for 8 hours
at a temperature of 100cC. After filtering and washing
with hot water, the residue assayed 0.5 percent arsenic.
2. Treatment of the Arsenic Pregnant Solution
The solution from the leaching step may contain, in
addition to the arsenic values, gold, antimony, mercury,
sodium and sulfur values which are recoverable and
which the process of this invention is designed to
recover. The arsenic removal and gold recovery is accomplished
in a series of steps which will now be
50 described.
A. Recovery of Gold Values
The ore concentrate described above contains approximately
1.55 ounces of gold concentrate. Although
leaching conditions may be controlled to regulate the
amount of gold that is dissolved, any gold present in the
arsenic pregnant solution may be removed by reducing
the pH of the solution to one within the range of from
about 9-11. This acidification step is preferably carried
out by contacting the solution with certain types of
acidifying agents as, for example, S02 gas. The contact
is carried out until the pH of the solution is at the
desired level, usually pH 10. At this point, any gold
present is precipitated, filtered, washed, dried for
recovery of the mineral values.
During the acidification step hydrogen sulfide is
evolved which is normally purified and recycled to the
process.
(approximate)
48.0%
19.0%
33.0%
PREPARATION OF CONCENTRATE
BRIEF DESCRIPTION OF THE INVENTION
1
PROCESS FOR REMOVAL OF ARSENIC FROM
SULFO·ORE
BRIEF STATEMENT OF THE INVENTION
The novel process of this invention may be logically
divided into a number of sequential steps as follows:
I. Leaching of the concentrate with a caustic solution;
2. Treatment of the arsenic pregnant solution; and
3. Treatment of the arsenic barren solution.
There are a number of processing steps involved in
each of the above-mentioned general classifications in
which conditions of the reaction are critical and must
be carefully controlled to obtain the optimum results.
These critical conditions will be particularly noted as
this description proceeds.
This invention relates to a process for the removal of
arsenic from materials in which it is contained. Particularly
the invention relates to a novel process for the
removal of arsenic from mixed ore mineral concentrates
containing a minor proportion thereof.
More particularly, the invention relates to an improved
process for the removal and/or recovery of arsenic
from ore concentrates which may contain mixed
values of copper-gold-sulfur-arsenic-antimony and
which involves the steps of leaching the arsenic values
from the ore concentrate, causing arsenic compounds
to precipitate from the solution and recovering values
such as gold, antimony, etc., from the leach solution.
Arsenic represents an undesirable contaminant in
smelting operations particularly in a number of coppercontaining
minerals. Exemplary of such minerals are
the sulfo-arsenic salts of copper, sulfo-antimony salts of
copper and mixtures such as tennantite, luzonite, 40
enargite, tetrahedrite, famantinite and the like. One
particularly valuable ore mineral containing components
in which the arsenic content of approximately
19 percent is especially undesirable is enargite having 45
approximately the following composition:
Copper
Arsenic
Sulfur
This particular mineral is often found to include
minor amounts of gold, silver, antimony, mercury,
bismuth. tin, zinc, lead and other metals.
Generally speaking, the concentrate is prepared by a 55
selective flotation operation, as is well known in the art.
I. Leaching of the Concentrate with a Caustic Solution
In accordance with the inventive process the concentrate
is leached with a hot concentrated caustic solution,
preferably of sodium sulfide, although it is possi- 60
ble to use sodium hydroxide or mixtures of sodium sulfide
and sodium hydroxide, or sodium hydroxide and
sulfur. Since the undissolved residue from the leaching
step contains valuable mineral valves, the undissolved 65
solids from the leaching step are settled, filtered,
washed and dried. They are then removed from the
process for metal recovery by conventional techniques.
3
3,709,680
4
It is to be understood that the apparatus which is
hereinafter described is only illustrative of one type of
apparatus which may be used to obtain the desired
results and the specification thereof is simply to
facilitate an understanding of the various steps involved.
This exemplary iIlustration is not to be considered
as a limitation of any kind on the apparatus as
defined in the appended claims.
Turning now to the drawings, and with particular
reference to FIG. 2, there is shown in the block diagram
form an arrangement of apparatus wherein the
feed to the process, an ore concentrate, is fed to
leaching vessel 10, which may be equipped with an
agitator driven by a motor through line 16. It will be appreciated,
of course, that for handling solid materials,
line 16 may be a screw or other type of conveyor means
and the latching vessel wiII be equipped with a hopper
as a feed inlet. If desired the leaching zone may be a
plurality of leaching vessels and may be provided with
means for temperature control as known to the art.
Hot sodium sulfide solution is added to leaching vessel
10 through line 18. After the desired leaching time
the slurry is passed to a series of thickening vessels arranged
for countercurrent washing, designated here by
numeral 20. Each thickener may be provided with a
scraper or rake to move the thickened slutty or residue
toward the center of the vessel bottom.
The leach solution is introduced to the thickening
zone by means of line 38. Water at the desired temperature
is introduced to the thickener 20, such that
the slurry of leaching solution and solids passes
counter-currently thereto.
Thickened solids are withdrawn from settler 20
35 through line 48 and passed to filter 50. The filtrate is
recycled to the process through line 52. Separated
solids, which represent the mineral values in the feed
concentrate, are removed from the filter through line
54 and set for further processing.
The counter-current wash or decant from the settlers
or thickening zone is withdrawn through line 56 and introduced
into first acidification vessel 62. This vessel,
which may be any of the known liquid-gas contacting
vessels known to the art, is equipped with inlet line 62
and a gas distribution means for bubbling an acidifying
gas, such as S02, through a column of the arsenic
pregnant solution in the vessel.
Acidification is continued in this vessel until a pH of
within a range of from about 9-11, preferably 10, is
reached. The solution with precipitated gold values,
normal1y in the form of its sulfide, is then withdrawn
through line 66, for example, and introduced to
thickener 68 which may be equipped with a rake driven
by a motor.
In this thickener the solids are allowed to settle,
gathered by the rake and withdrawn through line 74 to
filter 76. Separated solids which contain the
precipitated gold sulfides, are removed from the filter
through line 77 for processing to recover the gold con-
60 tent thereof.
The arsenic pregnant solution from thickener 68,
with gold content removed, and the filtrate from filter
76, through lines 78 and 80, is passed into a second
65 acidification vessel 82. In this liquid-gas contacting vessel,
as in vessel 62, an acidifying gas, such as S02is introduced
at the bottom, through line 86 and a sparger,
so that intimate liquid gas contact is obtained and
DETAILED DESCRIPTION OF THE APPARATUS
OF THE INVENTION
The invention is more clearly explained by reference
to the following drawings in which:
FIG. 1 represents a block diagram of the flow sheet
of one embodiment of the invention; and
FIG. 2 is a process block diagram of one form of apparatus
useful for carrying out the process.
B. Removal of Arsenic
The arsenic pregnant solution which has been
treated to recover dissolved gold is now subjected to a
second acidification step. Certain types of acidifying
agents are again used, such as S02 gas, and the solution 5
is contacted until the pH is reduced to one within a
range of about 2 to 6, preferably about 3 to 4. The H2S
generated by this acidification step is cycled for use at a
subsequent step in the process.
At a pH of 3 and at a temperature of 25° to 50°C, 10
preferably below about 50°C, the arsenic values in the
arsenic pregnant solution precipitate in the form of sulfides,
and the solid pulp is filtered and washed. The
solids from the filtration step comprise essentially the
arsenic sulfides and are normally discarded. These are 15
normally the tri- or pentasulfides.
3. Treatment of the Arsenic Barren Solution
The arsenic barren solution contains sodium and sulfur
values which are recovered and recycled for use.
In summary, this recovery involves evaporation of 20
the solution to crystalIize these values in the form of
salts, roasting the salts to recover sulfur in the form of
S02 gas and thereby oxidizing the sodium salts to
sulfates and thereafter reducing the calcine with a
source of carbon to convert the sodium values to sodi- 25
urn sulfide which is dissolved and recycled to the
leaching step.
A. Evaporation Step
Evaporation of the arsenic barren solution is carried
out in evaporation equipment known to the art and 30
yields a variety of sodium-sulfur compounds.
If desired a portion of the sulfur obtained during the
evaporation step may be removed prior to passing the
material to the oxidation roast.
B. The Oxidation Roast
The solids from the evaporation step containing the
sulfur and sodium values are conveyed to a roasting
zone where they are subjected to temperatures in the
order of 150° to 600°C, preferably 400° to 500°C. The
roasting is conducted in an oxidizing atmosphere nor- 40
mally in the form of air. H2S obtained from the previous
step may also be introduced. The effluent gases from
the roasting step contain approximately 13% S02 by
volume, which are recycled to the acidification step.
The calcine obtained from this roast is passed to the 45
next step in the process.
C. The Reducing Roast
The calcine from the first roasting step is preferably
admixed with a source of carbon, normally in the form
of a pulverized coal. After subjecting the calcine to an 50
adequate temperature for the desired reaction time, the
sodium sulfate contained in the calcine is reduced to
sodium sulfide.
Product from this second roasting or furnacing step
contains about 75% Na2S, the remainder being ash, un- 55
burned coal, and impurities. The product is then mixed
with water and the dissolved sodium sulfide is recycled
to the process.
3,709,680
7 8
furnacing said solid residue at a temperature of from
about 850° to I,200°C in the presence of carbon to
recover solid sodium sulfide therefrom;
dissolving solid sodium sulfide in water to form a
solution thereof; and 5
recycling sodium sulfide solution to the process.
* * * * *
10
15
20
25
30
35
40
45
50
55
60
65
-size:7f�G;o��0�:"Times New Roman","serif";mso-fareast-font-family: HiddenHorzOCR'>Molybdenum _
Potassium_ •• _
Silicon_ • _
Sodium • •
Tin • _
Others_ • _
I Not dried.
Tables 1 and 2 show that up to 98% of the molybdenum
contained in the original feed samples was recovered by
the process. Tables 1, 2, 5 and 6 show that up to 99%
of rhenium contained in the original samples was recovered
by the process. The final products obtained were
substantially free of impurities derived from the feed
solution.
The invention described provides an effective and
economical method for almost 1<00% recovery of molybdenum
and rhenium from solutions in which they are
present together. Although the solutions from which
the recoveries are made are ordinarily scrubber solutions
resulting from roasting of molybdenite concentrates, the
invention is not restricted to recovery of the metals from
this type solution.
What is claimed is:
1. A process for recovering molybdenum and rhenium
values from solutions in which they are present together
with other metal ion impurities which comprises:
(a) extracting the rhenium and molybdenum values
from the solution with a liquid water insoluble amine
ion exchange agent;
(b) stripping the rhenium and molybdenum values from
the loaded agent of (a) with a basic solution of an
ammonium compound;
(c) crystallizing molybdenum as ammonium tetramolybdate
from the ship solution of (b) by adjusting
the pH of the strip solution to about 2.0-3.5;
157 15
81
45
27
5.9
5.4
3.6o
51
28
14
8
1.1
0.6
0.5o
TABLE 5
Data for Extraction of Rhenium and Rejection of Molybdenum by Pyridine from 6 M Sodium Hydroxide
Solution
Molybdenum Rhenium
Grams per liter Grams per liter
Loaded Percent Loaded Percent
Test No. Feed solvent Raffinate extracted E, alo Feed solvent Raffinate extracted E,a!o
L __ e __•• __ 14.5 0.35 11.4 5 0.03 3.05 1.7 0.010 99.6 170
32_____________.:_c__. 17.1 0 18.4 0 0 1.42 1.2 0.008 99.5 155
14.5 0.09 9.9 5 0.009 3.05 0.40 0.003 99.8 133
4______ •• _. 14.5 0.94 10.8 20 0.09 3.05 3.3 0.03 99.0 110
5___ •• ____ • 14.5 0.07 10.5 3 0.007 3.05 0.84 0.01 99.5 84 6____ • ___ •• 18.4 ___ • __ • ___ • ___• _______ ••••c_._.__________ • 1.5 1.48 0.018 99.1 82 7_.________
40.9 0 41.1 0 0 16.8 8.7 0.107 99.3 81
8__ • _______ 18.4 • ____________• __ • _________________________
9____ • _____ 1.5 1.48 0.025 98.7 59
40.9 0 45.0 0 0 16.8 17.6 0.326 98.0 54
10_________ 40.9 0 46.1 0 0 16.8 44.7 5.26 69 8.5
Portions of rhenium have been recovered from the extract
by both of these procedures. In the first, the rhenium 70
was precipitated from the distillation bottoms as the
rhenium heptasulfide. In the second, the rhenium was reduced
to metal with hydrogen. The potassium and sodium
salts were removed by leaching with water and dilute hydrochloric
acid. Table 6 shows the quality of the rhenium 75
(d) recovering the crystallized ammonium tetramolybdate
of (c) followed by recovery of molybdenum
values therefrom;
(e) extracting rhenium 'Values from the mother liquor
of (d) with a liquid water insoluble amine ion exchange
agent;
3,681,016
25 HERBERT T. CARl1ER, Primary Examiner
u.s. Cl. X.R.
8
ing the pH of rthe solution' to a pH between about
2.5-3.5;
(d) separating the crystallized ammonium tetramolybdate
from the mother liquor of (c);
(e) extracting rhenium values from the mother liquor
of (d) with a liquid water insoluble amine ion exchange
agent;
(f) stripping rhenium values from the loaded agent of
(e) with sodium hydroxide;
(g) recovering rhenium values from the strip solution
of (f) by extracting with pyridine; and
(h) recovering rhenium from the pyridine extractant
by distilling off the pyridine.
References Cited
UNITED STATES PATENTS
7/10969 Litz , 23-15 W
7/1969 !Platzke et al. 23-15W
3/1959 Zimmerley et al. 23-18 X
7/1960 Zimmerley et al. 23-24
4/1966 Churchward , 23~15 W
211970 Ziegenbaly et al. 23-23 X
1/1971 Proter et al. 23-22
3,455,'677
3,45:8,277
2,876,065
2,945,743
3,244,475
3,495,934
3,558,268
23-23, 24 R, 51 R
7
(f) stripping the loaded agent of (e) with an alkali
metal hydroxide;
(g) extracting rhenium values from the strip solution
of (f) with pyridine or pyridine derivative; and
(h) recovering rhenium from the pyridine extractant 5
by distilling off the pyridine.
2. The process of claim 1 in which metal ion impurities
are removed from the strip solution of .(b) before
crystallizing ammonium tetramolybdate in (c).
3. The process of claim 1 in which the anion exchange 10
agent in (a) is a tertiary amine ion exchange resin and the
stripping solution of (b) is ammonium hydroxide.
4. A process for recovering molybdenum and rhenium
values from pregnant acid leach solutions containing these
values together with other metal impurities and derived 15
from dusts and flue gases resulting from roasting relatively
impure molybdenite concentrate, said process comprising:
(a) extracting molybdenum and rhenium values from
the pregnant acid solution with a liquid water in- 20
soluble amine ion exchange agent;
(b) stripping the molybdenum and rhenium values
from the exchange resin with ammonium hydroxide
solution to form a strip solution containing the molybdenum
as ammonium molybdate and the rhenium
as ammonium perrhenate;
(c) crystallizing the molybendum from the strip solution
in (b) as ammonium tetramolybdate by adjust