111111111111111111111111111111111111111111111111111111111111111111111111111
US007341700B2
(12) United States Patent
Marsden et al.
(10) Patent No.:
(45) Date of Patent:
US 7,341,700 B2
*Mar.ll,2008
(21) Appl. No.: 101756,574
(73) Assignee: Phelps Dodge Corporation, Phoenix,
AZ (US)
(54) METHOD FOR RECOVERY OF METALS
FROM METAL-CONTAINING MATERIALS
USING MEDIUM TEMPERATURE
PRESSURE LEACHING
(75) Inventors: John O. Marsden, Phoenix, AZ (US);
Robert E. Brewer, Safford, AZ (US);
Joanna M. Robertson, Thatcher, AZ
(US); Wayne W. Hazen, Lakewood,
CO (US); Philip Thompson, West
Valley City, UT (US); David R.
Baughman, Golden, CO (US); Roland
Schmidt, Golden, CO (US)
11/1976 Queneau et al.
4/1977 Johnson
4/1977 Ackerley et al.
6/1977 Domic et aI.
6/1977 Faugeras et al.
8/1977 Wong
8/1977 Stanley et al.
9/1977 Subramanian et aI.
1/1978 Wong
5/1978 Riggs et aI.
10/1978 Fountain et al.
4/1979 Dain
6/1979 Weir et al.
8/1979 Reynolds
3/1981 Baczek et al.
5/1981 Redondo-Abad et al.
6/1981 Lamb
7/1982 Stanley et al.
9/1983 Dienstbach
11/1983 Wilkomirsky et al.
4/1984 Baglin et al.
3/1985 Kordosky et al.
2/1986 Weir et al.
10/1986 Salter et aI.
10/1988 Horton et al.
3,991,159 A
4,017,309 A
4,020,106 A
4,028,462 A
4,029,733 A
4,039,405 A
4,039,406 A
4,046,851 A
4,069,119 A
4,091,070 A
4,120,935 A
4,150,976 A
4,157,912 A
4,165,362 A
4,256,553 A
4,266,972 A
4,272,341 A
4,338,168 A
4,405,569 A
4,415,540 A
4,442,072 A
4,507,268 A
4,571,264 A
4,619,814 A
Jan. 12, 2004 4,775,413 A
This patent is subject to a tenninal disclaimer.
Subject to any disclaimer, the tenn of this
patent is extended or adjusted under 35
U.S.c. 154(b) by 461 days.
( *) Notice:
(22) Filed:
(65) Prior Publication Data
12/1958
6/1999
4/2001
ABSTRACT
(Continued)
FOREIGN PATENT DOCUMENTS
OTHER PUBLICATIONS
0219785
1657-2000
WO 01/08890
(57)
Evans, et aI., "International Symposium of Hydrometallurgy," Mar.
1, 1973, 2 pages.
AU
CL
WO
(Continued)
Primary Examiner-Steven Bos
(74) Attorney, Agent, or Firm-Snell & Wilmer L.L.P.
US 2004/0146438 Al luI. 29, 2004
Related U.S. Application Data
(63) Continuation of application No. 09/915,105, filed on
luI. 25, 2001, now Pat. No. 6,676,909.
(60) Provisional application No. 60/220,673, filed on luI.
25,2000.
(51) Int. Cl.
C22B 15/00 (2006.01)
(52) U.S. Cl. 423/28; 423/20; 423/21.1;
423/22; 423/49; 423/53; 423/109; 423/150.1;
241/23; 75/743; 75/744
(58) Field of Classification Search 423/28,
423/21.5,22,49,53,20,109, 150.1, DIG. 15,
423/21.1; 241/23; 75/743, 744
See application file for complete search history.
(56) References Cited
U.S. PATENT DOCUMENTS
3,260,593 A
3,528,784 A
3,656,888 A
3,669,651 A
3,673,371 A
3,775,099 A
3,868,440 A
3,896,208 A
3,949,051 A
3,958,985 A
3,961,028 A
3,962,402 A
3,967,958 A
3,985,553 A
7/1966 Zimmerley et al.
9/1970 Green
4/1972 Barry et al.
6/1972 Spedden et al.
6/1972 Smith et al.
11/1973 Coffield et al.
2/1975 Lindblad et al.
7/1975 Dubeck et aI.
4/1976 Pawlek et al.
5/1976 Anderson
6/1976 Parker et al.
6/1976 Touro
7/1976 Coffield et al.
10/1976 Kunda et al.
The present invention relates generally to a process for
recovering copper and other metal values from metal-containing
materials using controlled, super-fine grinding and
medium temperature pressure leaching. Processes embodying
aspects of the present invention may be beneficial for
recovering a variety of metals such as copper, gold, silver,
nickel, cobalt, molybdenum, rhenium, zinc, uranium, and
platinum group metals, from metal-bearing materials, and
find particular utility in connection with the extraction of
copper from copper sulfide ores and concentrates.
15 Claims, 2 Drawing Sheets
u.s. PATENT DOCUMENTS
US 7,341,700 B2
Page 2
OTHER PUBLICATIONS
4,814,007 A
4,875,935 A
4,880,607 A
4,892,715 A
4,895,597 A
4,971,662 A
4,992,200 A
5,028,259 A
5,059,403 A
5,073,354 A
5,176,802 A
5,223,024 A
5,232,491 A *
5,316,567 A
5,356,457 A
5,431,717 A
5,573,575 A
5,645,708 A
5,650,057 A
5,670,035 A
5,676,733 A
5,698,170 A
5,730,776 A
5,770,170 A
5,849,172 A
5,869,012 A
5,874,055 A
5,895,633 A
5,902,474 A
5,914,441 A
5,917,116 A
5,985,221 A
5,989,311 A
5,993,635 A
6,083,730 A
6,146,444 A
6,149,883 A
6,428,604 Bl *
6,676,909 B2 *
3/1989 Lin et al.
10/1989 Gross et al.
11/1989 Horton et al.
111990 Horton
111990 Lin et al.
11/1990 Sawyer et al.
2/1991 Lin et al.
7/1991 Lin et al.
10/1991 Chen
12/1991 Fuller et al.
111993 Duyvesteyn et al.
6/1993 Jones
8/1993 Corrans et al 75/743
5/1994 Jones
10/1994 Alvarez et al.
7/1995 Kohr
11/1996 Kohr
7/1997 Jones
7/1997 Jones
9/1997 Virnig et al.
10/1997 Kohr
12/1997 King
3/1998 Collins et al.
6/1998 Collins et al.
12/1998 Allen et al.
2/1999 Jones
2/1999 Jones
4/1999 King
5/1999 Jones
6/1999 Hunter et al.
6/1999 Johnson et al.
11/1999 Knecht
11/1999 Han et al.
11/1999 Hourn et al.
7/2000 Kohr
11/2000 Kohr
1112000 Ketcham et al.
8/2002 Kerfoot et al. . 75/743
112004 Marsden et al 423/28
Duyesteyn, it aI., "The Escondida Process for Copper Concentrates,"
1998.
King, et aI., "The Total Pressure Oxidation of Copper Concentrates,"
1993.
King, J. A., "Autoclaving of Copper Concentrates," paper from
COPPER 95, vol. III: Electrorefining and Hydrometallurgy of
Copper, International Conference held in Santiago, Chile, Nov.
1995.
Mackis, V. N., "Direct Acid Pressure Leaching of Chalcocite
Concentrate," vol. 19, No.2, Feb. 1967.
Hirsch, H. E., "Leaching of Metal Sulphides," Patents, UK, No.
1,598,454, 1981, 7 pages.
Chimeilewski, T., "Pressure Leaching of a Sulphide Copper Concentrate
with Simultaneous Regeneration of the Leaching Agent,"
Hydrometallurgy, vol. 13, No.1, 1984, pp. 63-72.
Dannenberg, R. 0., "Recovery of Cobalt and Copper from Complex
Sulfide Concentrates," Government Report, 20 pages, Report No.
BM RI9138, U.S. Dept. of the Interior, 1987.
Berezowsky, R.M.G.S., "The Commercial Status ofPressure Leaching
Technology," JOM, vol. 43, No.2, 1991, pp. 9-15.
Hackl, R. P., "Effect of Sulfur-Dispersing Surfactants on the Oxygen
Pressure Leaching of Chalcopyrite," paper from COPPER 95, vol.
III, pp. 559-577, Met Soc of CIM, Nov. 1995.
Hackl, R. P., "Passivation of Chalcopyrite During Oxidative Leaching
in Sulfate Media," Hydrometallurgy, vol. 39, 1995, pp. 25-48.
L.W. Beckstead, et aI., "Acid Ferric Sulfate Leaching of AttritorGround
Chalcopyrite Concentrate," vol. II, Extractive Metallurgy of
Copper, Chapter 31, pp. 611-632, no date.
Jim A. King, et aI., paper entitled: "The Total Pressure Oxidation of
Copper Concentrates," vol. 1, Fundamental Aspects, 1993.
Dreisinger, D. B., "Total Pressure Oxidation of El Indio Ore and
Concentrate," COPPER 1999, Fourth International Conference,
Phoenix, Arizona, USA, Oct. 1999.
Richmond, G. D., "The Commissioning and Operation of a Copper
Sulphide Pressure Oxidation Leach Process at Mt. Grodon," ALTA
COPPER 1999: Copper Sulphides Symposium & Copper
Hydrometallurgy Fjorum, Gold Coast, Queensland, Australia Conference,
1999.
* cited by examiner
u.s. Patent Mar. 11,2008 Sheet 1 of 2 US 7,341,700 B2
102
104
106
108
110
METAL-BEARING MATERIAL
CONTROLLED,
SUPER-FINE GRINDING
PROCESSING
CONDITIONING
(OPTIONAL)
METAL RECOVERY
FIG. 1
- ---. ..... ~~ ...
100
98
*- 96
Z
0
t= 94 t:l
c:(
0= l- 92 xW
0w= 90
a..
a..
0 88
t:l
86
84
0
52
25 50 75
RESIDENCE TIME, MIN.
FIG. 3
100 125
u.s. Patent Mar. 11,2008 Sheet 2 of 2 US 7,341,700 B2
22 ; /20
V 24
206 """- CONTROLLED,
SUPER-FINE GRINDING
;28 V 26
30 V 27 31
/ 208.":-f I
PRESSURE LEACHING r--
V 32
210
ATMOSPHERIC FLASHING
V 34
58
212 """- SOLID-LIQUID J
PHASE SEPARATION
38
1/36 /
V 37
56
214
SOLVENT EXTRACTION /
V 40
42
216'j
SOLVENT STRIPPING J
V 44
48
218'-1 ; ELECTROLYTE RECYCLE TANK
V 48
54 52
\ 220"{ ; ELECTROWINNING I
V 50
( Cu )
FIG. 2
US 7,341,700 B2
1
METHOD FOR RECOVERY OF METALS
FROM METAL-CONTAINING MATERIALS
USING MEDIUM TEMPERATURE
PRESSURE LEACHING
CROSS-REFERENCE TO RELATED
APPLICATIONS
This application is a continuation of U.S. patent application,
Ser. No. 09/915,105, entitled "Method for Recovery of
Metals from Metal-Containing Materials Using Medium
Temperature Pressure Leaching," filed luI. 25, 2001, now
U.S. Patent 6,676,909, which claims priority to U.S. Provisional
Patent Application, Ser. No. 60/220,673 entitled
"Methods for Recovering Copper and Other Metals from
Sulfide Concentrates Using Medium Temperature Pressure
Oxidation," filed on luI. 25, 2000, both of which are
incorporated by reference herein.
FIELD OF INVENTION
The present invention relates generally to a process for
recovering copper and other metal values from metal-containing
materials, and more specifically, to a process for
recovering copper and other metal values from metal-containing
materials using controlled, super-fine grinding and
medium temperature pressure leaching.
BACKGROUND OF THE INVENTION
Smelting is a well-established approach for recovering a
metal, such as copper, from a metal-bearing sulfide material.
Due to the high cost of smelting, however, the copper sulfide
minerals in an ore body typically are first concentrated by
flotation techniques to provide a smaller volume for smelting.
The concentrate is then shipped to a smelter, which
processes the concentrate pyrometallurgically at high temperatures
to form a crude copper product that is subsequently
refined to a highly pure metal.
The recovery of copper from copper sulfide concentrates
using pressure leaching has proven to be a potentially
economically attractive alternative to smelting. Pressure
leaching operations generally produce less fugitive emissions
than smelting operations, and thus, enviroumental
benefits may be realized. Further, pressure leaching circuits
may be more cost-effectively constructed on-site at a concentrator,
eliminating the expense associated with concentrate
transportation that smelting operations may require.
Further, any byproduct acid produced in the pressure leaching
circuit may be used in adjacent heap leaching operations,
thus offsetting some of the costs associated with purchased
acid.
The mechanism by which pressure leaching processes
effectuate the release of copper from sulfide mineral matrices,
such as chalcopyrite, is generally dependent on temperature,
oxygen availability, and process chemistry. For
example, in high temperature pressure leaching processes
for chalcopyrite, that is, pressure leaching processes operating
above about 200° c., it has generally been found that
sulfur is fully converted to sulfate. In low temperature
pressure leaching processes (i.e., below about 100° C.), it
has generally been found that the chalcopyrite leaches
slowly and incompletely. Medium temperature pressure
leaching processes for chalcopyrite, which are generally
thought of as those processes operating at temperatures from
about 120° C. to about 190° c., have been the focus ofmuch
research and development in recent years and have shown
2
some promise for achieving a satisfactory compromise
between the high temperature and low temperature processes.
As discussed in further detail hereinbelow, however,
even with these efforts, such processes still exhibit significant
processing disadvantages.
Low temperature pressure leaching processes historically
have been disfavored because of characteristically low
extraction of copper and other metals, and long residence
times. High temperature pressure leaching processes, not-
10 withstanding their relatively short residence times and high
metal extractions, tend to have higher oxygen consumption,
higher by-product acid production, and greater heat production
in the pressure leaching vessel, which requires increased
cooling. Prior medium temperature pressure leaching pro-
15 cesses typically suffer incomplete copper extraction resulting
from either passivation of the copper sulfide particle
surfaces by a metal-polysulfide layer or partially-reacted
copper sulfide particles becoming coated with liquid
elemental sulfur and/or other reaction products. Further, in
20 prior medium temperature processes, under certain conditions,
molten elemental sulfur commonly agglomerates in
the pressure leaching vessel to form coarse sulfur "prills" or
"balls," which inhibit the extraction of copper and other
metals and which can create substantial difficulties with
25 materials handling and transport.
A variety of previous attempts have been made to circumvent
the problems associated with medium temperature
pressure leaching and to realize the potential benefits pursuant
thereto. For example, applying known pressure leach-
30 ing processes to the treatment of zinc sulfide materials,
previous attempts have been made to use surfactants such as
lignin derivatives, tannin compounds (such as quebracho),
and orthophenylene diamine (OPD) to disperse the elemental
sulfur formed and to render the copper in chalcopyrite
35 concentrates extractable. However, these attempts have not
been entirely successful since relatively low copper extraction
was realized even after significant residence times.
Other attempts have included pressure oxidation in the
presence of an acidic halide solution (U.S. Pat. No. 5,874,
40 055), and the use of finely divided particulate carbonaceous
material to inhibit passivation of incompletely leached copper
sulfide particles (U.S. Pat. No. 5,730,776). The feasibility
of using molten sulfur-dispersing surfactants to enhance
pressure leaching ofchalcopyrite in the temperature range of
45 125° C. to 155° C. has been investigated; however, it was
found that chalcopyrite particles (P90 of 25-38 microns)
leached too slowly even if molten sulfur was prevented from
passivating the material surfaces. See Hackl et aI., "Effect of
sulfur-dispersing surfactants on the oxidation pressure
50 leaching of chalcopyrite," proceedings of COPPER 95-COBRE
95 International Conference, Volume III, Electrorefining
and Hydrometallurgy of Copper, The Metallurgical
Society ofCIM, Montreal, Canada. The authors ofthat study
ultimately reported that the reaction rate for chalcopyrite
55 was controlled, at least in part, by a passivating mechanism
unrelated to sulfur formation.
It is generally known that hydrometallurgical processes,
particularly pressure leaching processes, are sensitive to
particle size. Thus, it is common practice in the area of
60 extractive hydrometallurgy to finely divide, grind, and/or
mill mineral species to reduce particle sizes prior to processing
by pressure leaching. For example, U.S. Pat. No.
5,232,491 to Corrans, et aI., entitled "Activation of a Mineral
Species," teaches a method of activating a mineral
65 species for oxidative hydrometallurgy by milling the species
to P80 of about 30 microns or less. Further, International
Publication No. WO 01/00890 to Anglo American PLC,
US 7,341,700 B2
3 4
entitled "Process for the Extraction of Copper," discusses
pressure leaching of copper sulfide particles (P80 from 5-20
microns) in the presence of a surfactant material at temperatures
from 1300 C. to 1600 C. According to test data set forth
in this publication, pressure leaching of chalcopyrite under
these conditions resulted in somewhat favorable copper
extractions ranging from about 88.2 to about 97.9%.
It generally has been appreciated that reducing the particle
size of a mineral species, such as, for example, copper
sulfide, enables pressure leaching under less extreme con- 10
ditions of pressure and temperature. The present inventors
have observed, however, that in addition to being sensitive
to the overall particle size distribution ofthe mineral species
being processed, pressure leaching processes-namely, copper
extraction by medium temperature pressure leaching 15
processes-are sensitive to the coarsest particle sizes in the
process stream above about 25 microns. Indeed, photomicrographs
of autoclave residue from coarse-ground (i.e., P80
of about 30-100 microns) chalcopyrite feed material have
indicated that unreacted chalcopyrite particles coarser than 20
about 20 microns were encapsulated in elemental sulfur. It
was observed that very few chalcopyrite particles finer than
about 10 microns remained in the residue.
An effective and efficient method to recover copper from
copper-containing materials, especially copper from copper 25
sulfides such as chalcopyrite and chalcocite, that enables
high copper recovery ratios at a reduced cost over conventional
processing techniques would be advantageous.
the invention, the consumption of acid is reduced, resulting
in a lower make-up acid requirement.
These and other advantages of a process according to
various aspects of the present invention will be apparent to
those skilled in the art upon reading and understanding the
following detailed description with reference to the accompanying
figures.
BRIEF DESCRIPTION OF THE DRAWINGS
The subject matter of the present invention is particularly
pointed out and distinctly claimed in the concluding portion
of the specification. A more complete understanding of the
present invention, however, may best be obtained by referring
to the detailed description and claims when considered
in counection with the drawing figures, wherein like numerals
denote like elements and wherein:
FIG. 1 illustrates a flow diagram of a copper recovery
process in accordance with the present invention;
FIG. 2 illustrates a flow diagram of a copper recovery
process in accordance with another embodiment of the
present invention; and,
FIG. 3 illustrates a graphical profile of copper extraction
as a function of temperature and time in accordance with
various embodiments of the present invention.
DETAILED DESCRIPTION
The present invention exhibits significant advancements
30 over prior art processes, particularly with regard to recovery
ratios and process efficiency. Moreover, existing copper
recovery processes that utilize a conventional atmospheric
or pressure leaching/solvent extraction/electrowinning process
sequence may, in many instances, be easily retrofitted
35 to exploit the many commercial benefits the present invention
provides.
Referring to FIG. 1, in accordance with various aspects of
the present invention, a metal-bearing material 102 is provided
for processing. Metal-bearing material 102 may be an
40 ore, a concentrate, or any other material from which copper
and/or other metal values may be recovered. Metal values
such as, for example, copper, gold, silver, zinc, platinum
group metals, nickel, cobalt, molybdenum, rhenium, uranium,
rare earth metals, and the like, may be recovered from
45 metal-bearing materials in accordance with various embodiments
of the present invention. The various aspects and
embodiments ofthe present invention, however, prove especially
advantageous in connection with the recovery of
copper from copper sulfide ores, such as, for example, ores
50 and/or concentrates containing chalcopyrite (CuFeS2), chalcocite
(Cu2S), bornite (Cu5FeS4), and covellite (CuS), and
mixtures thereof. Thus, metal-bearing material 102 preferably
is a copper ore or concentrate, and most preferably, is
a copper sulfide ore or concentrate.
Metal-bearing material 102 may be prepared for metal
recovery processing in any manner that enables the conditions
of metal-bearing material102-such as, for example,
composition and component concentration-to be suitable
for the chosen processing method, as such conditions may
60 affect the overall effectiveness and efficiency of processing
operations. Desired composition and component concentration
parameters can be achieved through a variety of chemical
and/or physical processing stages, the choice of which
will depend upon the operating parameters of the chosen
65 processing scheme, equipment cost and material specifications.
For example, as discussed in some detail hereinbelow,
metal-bearing material 102 may undergo comminution, flo-
SUMMARY OF THE INVENTION
While the way in which the present invention addresses
the deficiencies and disadvantages of the prior art is
described in greater detail hereinbelow, in general, according
to various aspects of the present invention, a process for
recovering copper and other metal values from a metalbearing
material includes various physical conditioning,
reactive, and recovery processes. In particular, controlled,
super-fine grinding of the metal-bearing material prior to
reactive processing enhances the recovery of copper and/or
other desired metal values. In accordance with the various
embodiments ofthe present invention, controlled, super-fine
grinding ofthe metal-bearing material prior to processing by
medium temperature pressure leaching results in enhanced
metal value recovery and various other advantages over
prior art metal recovery processes.
In accordance with an exemplary embodiment of the
present invention, a process for recovering copper from a
copper-containing material generally includes the steps of:
(i) providing a feed stream containing copper-containing
material; (ii) subjecting the copper-containing feed stream to
a controlled, super-fine grinding process; (iii) pressure
leaching the copper-containing feed stream to yield a copper-
containing solution; and (iv) recovering cathode copper
from the copper-containing solution. As used herein, the 55
term "pressure leaching" shall refer to a metal recovery
process in which material is contacted with an acidic solution
and oxygen under conditions of elevated temperature
and pressure. In one aspect of a preferred embodiment ofthe
invention, copper recovery of 98 percent is achievable while
still realizing various important economic benefits. In
another aspect of a preferred embodiment of the invention,
the use of a dispersing agent during pressure leaching
decreases undesirable agglomeration of elemental sulfur in
the pressure leaching vessel and passivation of unreacted
copper-bearing material particles by liquid elemental sulfur.
Moreover, in another aspect of a preferred embodiment of
US 7,341,700 B2
5 6
(2)
(1)
4CuFeS2+1702+4H20--;o2Fe203+4Cu2++8H++
8soi-
Preferably, in accordance with the present invention, the
temperature is suitably selected to achieve an advantageous
balance between reactions (1) and (2), but tending to reduce
acid consumption and thus the costs associated with acid
make-up, but without sacrificing copper extraction.
Reactive processing step 106 may occur in any pressure
leaching vessel suitably designed to contain the pressure
leaching mixture at the desired temperature and pressure
conditions for the requisite pressure leaching residence time.
In accordance with one aspect of a preferred embodiment of
the invention, the pressure leaching vessel used in process-
However, at high acid levels, copper extraction is lowered,
likely due to the wetting characteristics of the elemental
sulfur. At lower temperatures, acid is generally consumed
and elemental sulfur is formed according to the following
reaction:
2.4/4.8 mm Colorado sand, available from Oglebay Norton
Industrial Sands Inc., Colorado Springs, Colo. This silica
sand exhibits desirable characteristics such as roundness and
sphericity. However, any grinding medium that enables the
desired particle size distribution to be achieved may be used,
the type and size of which may be dependent upon the
application chosen, the product size desired, grinding apparatus
manufacturer's specifications, and the like. Exemplary
media include sand, silica, metal beads, ceramic beads, and
10 ceramic balls.
Preferably, grinding in accordance with the present invention
proceeds in a staged or closed-circuit manner. That is,
preferably the coarsest particles of metal-bearing material
102 are suitably ground to the desired level, while particles
15 already at or below the desired level are not subjected to
additional grinding. As such, cost savings can be obtained in
connection with grinding operations, while at the same time
limiting the size and weight proportion of the coarsest
particles.
Referring again to FIG. 1, after metal-bearing material
102 has been suitably prepared for processing by controlled,
super-fine grinding 104 and, optionally, other physical and/
or chemical conditioning processes, it is subjected to a
reactive processing step 106, for example, metal extraction.
25 However, reactive processing step 106 may be any suitable
process or reaction that puts the copper in metal-bearing
material 102 in a condition such that it may be subjected to
later copper recovery processing. In accordance with one
embodiment of the present invention, reactive processing
30 step 106 comprises medium temperature pressure leaching.
Preferably, reactive processing step 106 is a medium temperature
pressure leaching process operating at a temperature
in the range of about 1400 C. to about 1800 C. and more
preferably in the range of about 1500 C. to about 1750 C.
35 Generally, the present inventors have found that temperatures
above about 1600 c., and more preferably in the range
ofabout 1600 C. or about 1650 C. to about 1750 C. are useful
in connection with the various aspects of the present invention.
In accordance with a particularly preferred aspect of the
present invention, the optimum temperature range selected
for operation will tend to maximize the extraction of copper
and other metals, minimize acid consumption, and thereby
minimize make-up acid requirements. That is, at higher
45 temperatures, sulfide sulfur generally is converted to sulfate
according to the following reaction:
tation, blending, and/or slurry fonnation, as well as chemical
and/or physical conditioning before and/or after the controlled,
super-fine grinding stage.
In accordance with one aspect of the present invention,
metal-bearing material 102 is prepared for metal recovery
processing by controlled, super-fine grinding. Preferably, a
uniform, ultra-fine particle size distribution is achieved, as
experimental results suggest that copper extraction by
medium temperature pressure leaching is sensitive to the
coarsest sizes of copper-containing material particles in the
process stream. As discussed above, photomicrographs of
medium temperature pressure leaching residue from coarseground
chalcopyrite feed material (i.e., feed material not
subjected to controlled, super-fine grinding in accordance
with the present invention) have indicated that unreacted
chalcopyrite particles coarser than about 20 microns were
encapsulated in elemental sulfur. It was, however, observed
that very few chalcopyrite particles finer than about 10
microns remained in the residue. The present inventors have
achieved advancement in the art of copper hydrometallurgy 20
by recognizing that it is advantageous not only to reduce the
size of the copper-containing material particles in the process
stream, but also to ensure that the size and weight
proportion of the coarsest particles is minimized. Thus,
while the prior art generally teaches finely dividing, grinding,
and/or milling mineral species prior to extractive hydrometallurgical
processing such that, for example, approximately
80 percent of the particles are less than a certain size
(e.g., P80 of less than about 20 microns, see International
Publication No. WO 01/0890; P80 of less than about 30
microns, see u.s. Pat. No. 5,232,491; etc.), the prior art
generally allows a significant fraction (e.g., at least 20
percent) of the particles in the process stream to be larger
than about 20 microns. As mentioned above, particles
coarser than about 20 microns have been shown not to react
completely during medium temperature leaching, but are
occluded from reaction by elemental sulfur and/or other
byproducts. Significant advantages in processing efficiency
and copper recovery ratios are achievable by enabling
substantially all particles to react substantially completely. 40
For example, P80 distributions and other similar manners of
expressing size distributions do not generally enable such
results.
As used herein, the tenn "controlled, super-fine grinding"
refers to any process by which the particle size of the
material being processed is reduced such that substantially
all of the particles are small enough to react substantially
completely during medium temperature pressure leaching.
For example, in accordance with one aspect of the present
invention, a particle size distribution of approximately 98 50
percent passing about 25 microns is preferable, and more
preferably, the copper-containing material stream has a
particle size distribution of approximately 98 percent passing
from about 10 to about 23 microns, and optimally from
about 13 to about 15 microns. These particle size distribu- 55
tions were determined through the use of a Malvern optical
particle size analyzer. Other methods and apparatus, however,
may be utilized.
In accordance with one aspect of an exemplary embodiment
of the invention, satisfactory controlled, super-fine 60
grinding of chalcopyrite concentrate with an as-received
particle size of approximately 98 percent passing about 172
microns may be achieved using an Isamill ultra-fine grinding
apparatus, a stirred horizontal shaft ball mill with baffles
developed jointly by Mount Is a Mines (MIM), Australia, 65
and Netzsch Feinmahltechnik, Gennany. Preferably, if an
Isamill is utilized, the grinding media used is 1.2/2.4 mm or
US 7,341,700 B2
7 8
the like. A variety of factors, such as the process material
balance, environmental regulations, residue composition,
economic considerations, and the like, may affect the decision
whether to employ a CCD circuit, a thickener, a filter,
or any other suitable device in a solid-liquid separation
apparatus. However, it should be appreciated that any technique
of conditioning the product slurry for later metal value
recovery is within the scope of the present invention.
As further discussed hereinbelow, the separated solids
10 may further be subjected to later processing steps, including
precious metal or other metal value recovery, such as, for
example, recovery of gold, silver, platinum group metals,
molybdenum, zinc, nickel, cobalt, uranium, rhenium, rare
earth metals, and the like, by cyanidation or other tech-
15 niques. Alternatively, the separated solids may be subject to
impoundment or disposal.
The liquid separated from a solid-liquid separation apparatus
also may undergo a series of conditioning steps to
prepare the copper solubilized therein for recovery. For
20 example, the separated liquid may undergo various reagent
additions and/or solvent extraction stages to put the copper
in a state such that the copper is susceptible to conventional
copper recovery techniques. Further, subsequent conditioning
and/or processing steps may be undertaken such that
25 recovery rates are as efficient as possible.
After any desired preparation steps, the pressure leaching
product stream undergoes the desired copper recovery step.
The copper recovery step may include any suitable conditioning
and/or copper recovery method or methods, for
30 example, electrowinning, precipitation, solvent extraction
(sometimes referred to as solution extraction or liquid ion
exchange), ion exchange, and/or ion flotation, and preferably
results in a relatively pure copper product.
In an exemplary embodiment of the present invention
35 illustrated in FIG. 2, a copper-containing feed stream 20
containing a copper-bearing material is provided for metal
value recovery. The copper in the copper-bearing material
may be in any form from which copper may be extracted,
such as copper oxide or copper sulfide, for example chal-
40 copyrite (CuFeS2), chalcocite (Cu2S), bornite (Cu5FeS4),
and covellite (CuS). The copper-bearing material also may
include any number of a variety of other metals, such as
gold, silver, platinum group metals, zinc, nickel, cobalt,
molybdenum, rhenium, rare earth metals, uranium, and/or
45 mixtures thereof.
The feed stream of copper-bearing material can be provided
in any number of ways, such that the conditions ofthe
feed stream are suitable for the medium temperature pressure
leaching aspect of the present invention. For example,
50 feed stream conditions such as particle size, composition,
and component concentrations can affect the overall effectiveness
and efficiency of medium temperature pressure
leaching.
In accordance with one aspect of the invention, the initial
55 copper-bearing feed material may be comminuted to facilitate
fluid transport and/or to optimize the inlet conditions for
the controlled, super-fine grinding operation. A variety of
acceptable techniques and devices for reducing the particle
size of the copper-bearing material are currently available,
60 such as ball mills, tower mills, grinding mills, attrition mills,
stirred mills, horizontal mills and the like, and additional
techniques may later be developed that may achieve the
desired result of reducing the, particle size of the copperbearing
material to be transported.
FIG. 2 illustrates an embodiment of the present invention
wherein a copper-bearing material stream 24 is a copper
sulfide concentrate, such as a chalcopyrite concentrate. In
ing step 106 is an agitated, multi-compartment pressure
leaching vessel. However, it should be appreciated that any
pressure leaching vessel that suitably permits metal-bearing
material 102 to be prepared for copper recovery may be
utilized within the scope of the present invention.
During reactive processing step 106, copper and/or other
metal values may be solubilized or otherwise liberated in
preparation for later recovery processes. Any substance that
assists in solubilizing the metal value, and thus releasing the
metal value from a metal-bearing material, may be used. For
example, where copper is the metal being recovered, an acid,
such as sulfuric acid, may be contacted with the copperbearing
material such that the copper may be solubilized for
later recovery steps. However, it should be appreciated that
any suitable method of solubilizing metal values in preparation
for later metal recovery steps may be utilized within
the scope of this invention.
Subsequent to metal-bearing material 102 undergoing
reactive processing step 106, the copper and/or other metal
values that have been made available by the reactive process
undergo one or more of various metal recovery processes.
Referring again to FIG. 1, metal recovery process 110 may
be any process for recovering copper and/or other metal
values, and may include any number of preparatory or
conditioning steps (optional step 108). For example, a
copper-bearing solution may be prepared and conditioned
for metal recovery through one or more chemical and/or
physical processing steps. The product stream from reactive
processing step 106 may be conditioned to adjust the composition,
component concentrations, solids content, volume,
temperature, pressure, and/or other physical and/or chemical
parameters to desired values and thus to form a suitable
copper-bearing solution. Generally, a properly conditioned
copper-bearing solution will contain a relatively high concentration
of soluble copper, for example, copper sulfate, in
an acid solution and preferably will contain few impurities.
Moreover, the conditions of the copper-bearing solution
preferably are kept substantially constant to enhance the
quality and uniformity of the copper product ultimately
recovered.
In one aspect of a preferred embodiment of the present
invention, conditioning of a copper-containing solution for
copper recovery in an electrowinning circuit begins by
adjusting certain physical parameters of the product slurry
from the reactive processing step. In a preferred aspect of
this embodiment of the invention, wherein the reactive
processing step is medium temperature pressure leaching, it
is desirable to reduce the temperature and pressure of the
product slurry to approximately ambient conditions. A preferred
method of so adjusting the temperature and pressure
characteristics of the copper-containing product slurry from
a medium temperature pressure leaching stage is atmospheric
flashing. Further, flashed gases, solids, solutions, and
steam may optionally be suitably treated, for example, by
use of a venturi scrubber wherein water may be recovered
and hazardous materials may be prevented from entering the
enviroument. In accordance with further aspects of this
preferred embodiment, after the product slurry has been
subjected to atmospheric flashing using, for example, a flash
tank, to achieve approximately ambient conditions of pressure
and temperature, the product slurry may be further
conditioned in preparation for later metal-value recovery
steps. For example, one or more solid-liquid phase separation
stages may be used to separate solubilized metal solution
from solid particles. This may be accomplished in any 65
conventional mauner, including use of filtration systems,
counter-current decantation (CCD) circuits, thickeners, and
US 7,341,700 B2
9 10
this embodiment, inlet stream 27 has a preferred solid-liquid
ratio ranging from about 5 percent to about 50 percent solids
by weight, and preferably from about 10 percent to about 35
percent solids by weight.
Any agent capable of assisting in the solubilization of the
copper, such as, for example, sulfuric acid, may be provided
during the pressure leaching process in a number of ways.
For example, such acid may be provided in a cooling stream
provided by the recycle of the raffinate solution 56 from the
10 solvent extraction step 214 and/or by the production during
pressure leaching of a sulfuric acid from the oxidation of the
sulfide minerals in the feed slurry. However, it should be
appreciated that any method of providing for the solubilization
of copper is within the scope ofthe present invention.
15 The amount of acid added during pressure leaching preferably
is balanced according to the acid needed to optimize
copper extraction. When optimal copper recovery is
attained, the elemental sulfur formed as a reaction byproduct
becomes intimately associated with the hematite byproduct
20 as it is precipitated and generally does not significantly
impact the copper leaching reaction. At high (i.e., much
greater than stoichiometric) acid dosages, however, the
amount of hematite precipitated in the pressure leaching
vessel generally decreases and the byproduct elemental
25 sulfur may encapsulate and/or passivate unreacted chalcopyrite
particles. In addition, the sulfur may form agglomerates.
The formation of these elemental sulfur agglomerates---or
sulfur "prills" as they are sometimes called-is generally
associated with decreased copper recovery, as discussed
30 above.
The amount of acid introduced into medium temperature
pressure leaching vessel 208 varies depending upon the
reaction conditions, particularly, reaction temperature. In
certain cases, make-up acid is introduced on the order of
35 from about 300 to about 650 kilograms per tonne of concentrate,
or less; however, lower make-up acid is required at
higher temperatures. For example, at 1600 c., copper extraction
of 98.0% was achieved at a net chemical acid consumption
of320 kg/tonne. At 1700 c., copper extraction of98.0%
40 was achieved at a net chemical acid consumption of 250
kg/tonne. At 1800 c., copper extraction of 98.1% was
achieved at a net chemical acid consumption of 225
kg/tonne (however, during this test, prills may have been
formed and, as such, actual copper extraction may vary).
The medium temperature pressure leaching process in
pressure leaching vessel 208 occurs in a manner suitably
designed to promote substantially complete solubilization of
the copper. Various parameters influence the medium temperature
pressure leaching process. For example, during
50 pressure leaching, it may be desirable to introduce materials
to enhance the pressure leaching process. In accordance with
one aspect ofthe present invention, during pressure leaching
in pressure leaching vessel 208, sufficient oxygen 31 is
injected into the vessel to maintain an oxygen partial pres-
55 sure from about 50 to about 200 psi, preferably from about
75 to about 150 psi, and most preferably from about 100 to
about 125 psi. Furthermore, due to the nature of medium
temperature pressure leaching, the total operating pressure
in pressure leaching vessel 208 is generally superatmo-
60 spheric, preferably from about 100 to about 750 psi, more
preferably from about 300 to about 700 psi, and most
preferably from about 400 to about 600 psi.
The residence time for the medium temperature pressure
leaching process can vary, depending on factors such as, for
65 example, the characteristics of the copper-bearing material
and the operating pressure and temperature of the pressure
leaching vessel. In one aspect of a preferred embodiment of
one aspect of a preferred embodiment of the present invention,
the copper-bearing material stream 24 is fed from a
surge pile or tank (not shown) to a controlled, super-fine
grinding unit 206. Process water 22 is preferably added to
copper-bearing material stream 24 to bring the percent solids
to the optimal pulp density specified for the controlled,
super-fine grinding unit 206. In preparation for pressure
leaching processing (step 208), the particle size of copperbearing
material stream 24 is reduced in a controlled,
super-fine grinding unit 206. Controlled, super-fine grinding
unit 206 may comprise any milling or grinding apparatus or
combination of apparatus suitable to produce a fine, particle
size distribution for ground copper-containing material
stream 26. A variety of apparatus are available for this
purpose, including, for example, ball mills, tower mills,
attrition mills, stirred mills, horizontal mills, and the like,
and additional techniques and apparatus may later be developed
that may achieve the controlled, super-fine grinding
described herein. As previously mentioned, grinding in
accordance with the present invention may proceed in a
staged or closed-circuit manner. That is, preferably the
coarsest particles of metal-bearing material 102 are suitably
ground to a desired level, while particles already at the
desired level are not subjected to further grinding.
Controlled, super-fine grinding serves several functions
advantageous to the hydrometallurgical processing of copper
sulfides, such as chalcopyrite. First, it increases the
surface area of the copper sulfide particles, thereby increasing
reaction kinetics. Moreover, controlled, super-fine grinding
increases the liberation of copper sulfide mineral particles
from gangue and it reduces copper sulfide slurry
abrasion such that the slurry may be more easily introduced
to the pressure leaching unit. In accordance with one aspect
of the present invention, the particle size of the coppercontaining
material stream is reduced by controlled, superfine
grinding to a 98 percent passing size (i.e., P98) of less
than about 25 microns, and more preferably, to a P98 of from
about 10 to about 23 microns, and most preferably from
about 13 to about 15 microns.
In one aspect of a preferred embodiment of the present
invention, the controlled, super-finely ground copper-containing
material 26 is combined with a liquid 28 to form a
copper-containing inlet stream 27. Preferably, the liquid
comprises process water, but any suitable liquid may be
employed, such as, for example, recycled raffinate, pregnant 45
leach solution, or lean electrolyte.
The combination of liquid 28 with the controlled, superfinely
ground copper-containing material 26 can be effectuated
using anyone or more of a variety of techniques and
apparatus, such as, for example, in-line blending or using a
mixing tank or other suitable vessel. In accordance with a
preferred aspect of this embodiment, the material stream is
concentrated with the copper-containing material being on
the order less than about 50 percent by weight of the stream,
and preferably about 40 percent by weight of the stream.
Other concentrations that are suitable for transport and
subsequent processing may, however, be used.
With continued reference to FIG. 2, inlet stream 27 is
suitably introduced to a pressure leaching vessel to undergo
medium temperature pressure leaching; as such, the pressure
leaching vessel preferably comprises an agitated, multicompartment
pressure leaching vessel 208. As discussed in
detail above, inlet stream 27 preferably has a solid particle
size suitably dimensioned such that the size distribution of
no more than about 2% of the concentrated copper-containing
material is larger than about 23 microns (i.e., P98 ofless
than about 23 mm). In accordance with a preferred aspect of
11
US 7,341,700 B2
12
the invention, the residence time for the medium temperature
pressure leaching of chalcopyrite ranges from about 30
to about 180 minutes, more preferably from about 60 to
about 120 minutes, and most preferably on the order of
about 90 minutes.
Control of the pressure leaching process, including control
of the temperature in pressure leaching vessel 208, may
be accomplished by any conventional or hereafter devised
method. For example, with respect to temperature control,
preferably the pressure leaching vessel includes a feedback
temperature control feature. For example, in accordance
with one aspect of the invention, the temperature of the
pressure leaching vessel 208 is maintained at a temperature
in the range of about 140° C. to about 180° C. and more
preferably in the range of about 150° C. to about 175° C.
Generally, the present inventors have found that temperatures
above about 160° c., and more preferably in the range
ofabout 160° C. or about 165° C. to about 175° C. are useful
in connection with the various aspects of the present invention.
Due to the exothermic nature of pressure leaching of
metal sulfides, the heat generated by medium temperature
pressure leaching is generally more than that needed to heat
feed slurry 27 to the desired operating temperature. Thus, in
order to maintain preferable pressure leaching temperature,
a cooling liquid may be introduced into the pressure leaching
vessel during pressure leaching. In accordance with one
aspect ofthis embodiment ofthe present invention, a cooling
liquid is preferably contacted with the feed stream in pressure
leaching vessel 208 during pressure leaching. Cooling
liquid may comprise make-up water, but can be any suitable
cooling fluid from within the refining process or from an
outside source, such as recycled liquid phase from the
product slurry or a mixture of cooling fluids. Cooling liquid
may be introduced into pressure leaching vessel 208 through
the same inlet as feed slurry, or alternatively in any manner
that effectuates cooling of feed slurry 27. The amount of
cooling liquid added to feed slurry 27 during pressure
leaching may vary according to the amount of sulfide
minerals in and the pulp density ofthe feed slurry 27, as well
as other parameters of the pressure leaching process. In a
preferred aspect of this embodiment of the invention, a
sufficient amount of cooling liquid is added to pressure
leaching vessel 208 to yield a solids content in product slurry
32 on the order of less than about 50% solids by weight,
more preferably ranging from about 3 to about 35% solids
by weight, and most preferably ranging from about 8 to
about 20% solids by weight.
In accordance with a preferred aspect of the present
invention, medium temperature pressure leaching 208 of
inlet stream 27 is performed in the presence of a dispersing
agent 30. Suitable dispersing agents useful in accordance
with this aspect of the present invention include, for
example, organic compounds such as lignin derivatives,
such as, for example, calcium and sodium lignosulfonates,
tannin compounds, such as, for example, quebracho, orthophenylene
diamine (OPD), alkyl sulfonates, such as, for
example, sodium alkylbenzene sulfonates, and combinations
of the above. Dispersing agent 30 may be any compound
that resists substantial degradation in the temperature range
of medium temperature pressure leaching (i.e., from about
140° C. to about 180° C.) and that achieves the desired result
of preventing elemental sulfur produced during the medium
temperature pressure leaching process-and thus present in
the pressure leaching vessel-from agglomerating and from
wetting the surface of the copper-containing material being
processed. Dispersing agent 30 may be introduced to pressure
leaching vessel 208 in an amount and/or at a concentration
sufficient to achieve the desired result. In one aspect
ofa preferred embodiment ofthe invention, favorable results
are achievable during pressure leaching of chalcopyrite
using calcium lignosulfonate in an amount of about 2 to
about 20 kilograms per tonne, and more preferably in an
amount of about 10 kilograms per tonne of chalcopyrite
concentrate.
In accordance with a preferred aspect of the embodiment
ofthe invention illustrated in FIG. 2, product slurry 32 from
10 pressure leaching vessel 208 may be flashed in an atmospheric
flash tank 210 or other suitable vessel to release
pressure and to evaporatively cool product slurry 32 through
the release of steam to form a flashed product slurry 34.
Depending upon the specific process equipment configura-
15 tions and specifications, more than one flash stage may be
employed. Flashed product slurry 34 preferably has a temperature
ranging from about 90° C. to about 105° c., a
copper concentration of from about 35 to about 60 grams/
liter, and an acid concentration of from about 10 to about 60
20 grams/liter.
Referring still to FIG. 2, flashed product slurry 34 may be
directed to a solid-liquid separation apparatus 212, such as
a CCD circuit. Alternatively, the solid-liquid separation
apparatus may comprise, for example, a thickener or a filter.
25 In one aspect of a preferred embodiment of the invention,
solid-liquid phase separation step 212 may be carried out
with a conventional CCD utilizing conventional countercurrent
washing ofthe residue stream to recover leached copper
to the copper-containing solution product and to minimize
30 the amount of soluble copper advancing to precious metal
recovery processes or storage. Preferably, large wash ratios
are utilized to enhance the effectiveness of the solid-liquid
separation stage-that is, relatively large amounts of wash
water are added to the residue stream in CCD circuit 212.
35 Preferably, flash product slurry 34 is diluted by the wash
water in CCD circuit 212 to form a copper-containing
solution having a copper concentration of from about 30 to
about 60 grams/liter.
Depending on its composition, residue stream 58 from
40 solid-liquid separation apparatus 212 may be disposed of or
subjected to further processing, such as, for example, precious
metal recovery. For example, if residue stream 58
contains an economically significant fraction of gold, it may
be desirable to recover this gold fraction through a cyani-
45 dation process or other suitable recovery process. If gold or
other precious metals are to be recovered from residue
stream 58 by cyanidation techniques, the content of contaminants
in the stream, such as elemental sulfur, iron
precipitates, and unreacted copper minerals, is preferably
50 minimized. Such materials generally promote high reagent
consumption in the cyanidation process and thus increase
the expense of the precious metal recovery operation. Additionally,
as mentioned above, it is preferable to use a large
amount of wash water or other diluent during the solid-
55 liquid separation process to maintain low copper and acid
levels in the CCD residue in an attempt to optimize the
residue stream conditions for precious metal recovery.
Referring still to FIG. 2, in accordance with various
aspects ofthe present invention, the recovery of copper may
60 be accomplished through conventional solvent extraction
and electrowinning techniques. For example, a diluting
solution 38 may be contacted with the separated liquid 36
from solid-liquid separation apparatus 212 to reduce the acid
concentration of the separated liquid 36 sufficiently to
65 provide desirable equilibrium conditions for solvent extraction
214. Solution 38 may be any suitable liquid, for
example, water or atmospheric leach eflluent solution, that
US 7,341,700 B2
13 14
EXAMPLE I
As discussed in detail hereinabove, controlled, super-fine
grinding of chalcopyrite concentrates is preferred prior to
medium temperature pressure leaching at about 1400 C. to
about 1800 C. to prevent encapsulation ofunreacted copper
minerals by elemental sulfur and/or copper polysulfide. The
various grinding systems set forth below were used to
produce an ultra-finely ground inlet stream of chalcopyrite
concentrate samples containing approximately 30.5 percent
copper for a medium temperature pressure leaching pilot
plant. The as-received particle size of the chalcopyrite
concentrate sample used in the continuous pilot plant tests
was P98=approximately 101 microns. The as-received particle
size of the chalcopyrite concentrate sample used in the
batch tests was P98=approximately 172 microns.
1) Conventional regrind mill followed by a short grind in
a Union Process stirred pin mill-material was reground in
a conventional regrind mill for sixty (60) minutes followed
by five (5) minutes in a Union Process batch stirred mill.
2) Conventional regrind mill followed by a longer grind
in a Union Process stirred pin mill-material was reground
in a conventional regrind mill for sixty (60) minutes followed
by twenty (20) minutes in a Union Process batch
stirred mill.
3) Open circuit Metprotech mill-material was ground for
thirty (30) minutes in a continuous Metprotech vertical
stirred pin mill. Steel media (approximately 4 mm) was
used.
4) Closed circuit Metprotech mill-material was ground
for thirty (30) minutes in a continuous Metprotech mill, then
cycloned with a 2 inch cyclone. Underflow was ground for
fifteen (15) minutes in a continuous Metprotech mill and
combined with the cyclone overflow as final product.
5) Single pass Netzsch mill-material was ground in a
single pass using a continuous Netzsch 4 liter mill and a net
proper conditioning of the copper-containing solution, a
high quality, uniformly-plated cathode copper product may
be realized without subjecting the copper-containing solution
to solvent extraction prior to entering the electrowinning
circuit is within the scope of the present invention. As
previously noted, careful control of the conditions of the
copper-containing solution entering an electrowinning circuit---
especially maintenance of a substantially constant
copper composition in the stream-can enhance the quality
10 of the electrowon copper by, among other things, enabling
even plating of copper on the cathode and avoidance of
surface porosity in the cathode copper, which degrades the
copper product and thus diminishes its economic value. In
accordance with this aspect of the invention, such process
control can be accomplished using any of a variety of
techniques and equipment configurations, so long as the
chosen system and/or method maintains a sufficiently constant
feed stream to the electrowinning circuit. As those
skilled in the art are aware, a variety of methods and
apparatus are available for the electrowinning of copper and
other metal values, any of which may be suitable for use in
accordance with the present invention, provided the requisite
process parameters for the chosen method or apparatus
are satisfied. The Examples set forth hereinbelow are illustrative
of various aspects of certain preferred embodiments
of the present invention. The process conditions and parameters
reflected therein are intended to exemplify various
aspects of the invention, and are not intended to limit the
scope of the claimed invention.
sufficiently reduces the copper and acid concentrations to
desired levels. In a preferred aspect of this embodiment of
the invention, sufficient amount of solution 38 is contacted
with the separated liquid stream 36 to yield an acid concentration
in the diluted copper-containing solution 37 preferably
ranging from about 2 to about 25 grams/liter, and more
preferably from about 4 to about 7 grams/liter and a pH
preferably ranging from about pH 1.5 to about pH 2.5 and
more preferably from about pH 1.8 to about pH 2.2, and
optimally in the range of about pH 2.0.
The diluted copper-containing solution 37 may be further
processed in a solvent extraction step 214. During solvent
extraction 214, copper from copper-containing solution 29
may be loaded selectively onto an organic chelating agent,
for example, an aldoxime/ketoxime blend, resulting in a 15
copper-containing organic stream 40 and a raffinate solution
56. Raffinate 56 from solvent extraction step 214 may be
used beneficially in a number of ways. For example, all or
a portion of raffinate 56 maybe recycled to pressure leaching
vessel 10 for temperature control or may be used in heap 20
leaching operations, or may be used for a combination
thereof. The use of raffinate 56 in heap leaching operations
may be beneficial because the acid and ferric iron values
contained in raffinate 56 can act to optimize the potential for
leaching oxide and/or sulfide ores that commonly dominate 25
heap leaching operations. That is, the ferric and acid concentrations
of raffinate 56 may be used to optimize the Eh
and pH ofheap leaching operations. It should be appreciated
that the properties of raffinate 56, such as component concentrations,
may be adjusted in accordance with the desired 30
use of raffinate 56.
Copper-containing organic stream 40 is then subjected to
a solvent stripping phase 216, wherein more acidic conditions
are used to shift the equilibrium conditions to cause the
copper in the reagents to be exchanged for the acid in a 35
highly acidic stripping solution. As shown in FIG. 2, an
acid-bearing reagent 42, preferably sulfuric acid, and optionally,
lean electrolyte 54, are contacted with copper-containing
organic stream 40 during solvent stripping phase 216.
Sulfuric acid is a preferred acid-bearing reagent and is a 40
desirable copper matrix for electrowinning operations. The
acid-bearing reagent is contacted with the copper-containing
organic stream to effectuate the exchange of acid for copper
to provide copper for electrowinning.
Referring still to FIG. 2, copper-containing solution 45
stream 44 from solvent stripping phase 216 may be sent to
an electrolyte recycle tank 218. The electrolyte recycle tank
may suitably facilitate process control for electrowinning
stage 220, as will be discussed in greater detail below.
Copper-containing solution stream 44, which generally con- 50
tains from about 35 to about 50 grams/liter of copper and
from about 145 to about 180 grams/liter acid, is preferably
blended with a lean electrolyte 54 (i.e., electrolyte that has
already been through the metal recovery phase and has had
a portion of its dissolved copper removed) and makeup fluid 55
46, such as, for example, water, in the electrolyte recycle
tank 218 at a ratio suitable to yield a product stream 48, the
conditions ofwhich may be chosen to optimize the resultant
product of electrowinning step 220.
Preferably, the copper composition of product stream 48 60
is maintained substantially constant at a value from about 20
to about 60 grams/liter, more preferably at a value from
about 30 to about 50 grams/liter. Copper values from the
copper-containing product stream 48 are removed during
electrowinning step 220 to yield a pure, cathode copper 65
product. It should be appreciated that in accordance with the
various aspects of the invention, a process wherein, upon
US 7,341,700 B2
16
EXAMPLE 3
FIG. 3 is a graphical profile of continuous pilot plant test
data illustrating copper extraction as a function of time in
accordance with various embodiments of the present invention.
For each test run, the chalcopyrite concentrate samples
were ground to a P98 ofless than about 23 microns. Calcium
lignosulfonate from Georgia Pacific Corp. was introduced to
the pressure leaching vessels in an amount of about 10
kilograms per tonne of concentrate.
Curve 52 illustrates copper extraction versus residence
time for medium temperature pressure leaching of chalcopyrite
at approximately 1600 c., with acid addition to the
pressure leaching vessel of about 580 kilograms per tonne.
Approximately 96% copper extraction was achieved at
about 60 minutes, and 98+% copper extraction was achieved
at a residence time of about 95 minutes.
Curve 54 illustrates copper extraction versus residence
time for medium temperature pressure leaching of chalcopy-
20 rite at approximately 1700 c., with acid addition to the
pressure leaching vessel of about 507 kilograms per tonne.
Approximately 96% copper extraction was achieved at
about 60 minutes, and 98+% copper extraction was achieved
at a residence time of about 80 minutes.
Curve 56 illustrates copper extraction versus residence
time for medium temperature pressure leaching of chalcopyrite
at approximately 1800 c., with acid addition to the
pressure leaching vessel of about 421 kilograms per tonne.
Approximately 96% copper extraction was achieved at
about 52 minutes, and 98+% copper extraction was achieved
at a residence time of about 90 minutes (however during this
test prills may have been fonned and, as such, actual copper
extraction may vary).
An effective and efficient method to recover copper from
copper-containing materials, especially copper from copper
sulfides, such as chalcopyrite, that enables high copper
recovery ratios at a reduced cost over conventional processing
techniques has been presented herein. In accordance
with the present invention, it has been shown that copper
recovery in excess of 98 percent is achievable while realizing
various important economic benefits of medium temperature
pressure leaching and circumventing processing
problems historically associated with medium temperature
pressure leaching. The use of a dispersing agent during
pressure leaching lessens undesirable agglomeration of
elemental sulfur in the pressure leaching vessel and passivation
of unreacted copper-bearing material particles by
liquid elemental sulfur. Further, the present inventors
advanced the art of copper hydrometallurgy by recognizing
the advantages of not only reducing the size of the coppercontaining
material particles in the process stream, but also
ensuring that the size and weight proportion of the coarsest
particles are minimized.
The present invention has been described above with
reference to a number of exemplary embodiments and
examples. It should be appreciated that the particular
embodiments shown and described herein are illustrative of
the invention and its best mode and are not intended to limit
in any way the scope of the invention as set forth in the
claims. Those skilled in the art having read this disclosure
will recognize that changes and modifications may be made
to the exemplary embodiments without departing from the
scope of the present invention. Further, although certain
preferred aspects of the invention are described herein in
tenns of exemplary embodiments, such aspects ofthe invention
may be achieved through any number of suitable means
now known or hereafter devised. Accordingly, these and
15
TABLE I
Copper Extraction versus Grind Fineness in Continuous
Pilot Plant Tests
energy input of 56 kWhr/tonne. Colorado sand media (1.21
2.4 mm or 2.414.8 mm) was used as the grinding media.
6) Double pass Netzsch mill-material was ground twice
in the continuous Netzsch mill. The single pass material was
ground in another pass through the mill using a net energy
input of 56 kWhr/tonne for the second pass. Colorado sand
media (1.2/2.4 mm) was used as the grinding media.
Continuous pilot plant results indicate that copper extraction
is sensitive to grind fineness. For example, it was 10
observed that a grind fineness of approximately 98 percent
passing about 23 microns was required to achieve approximately
98 percent copper extraction at about 1600 C. and
about 500 kgltonne sulfuric acid addition to the pressure 15
leaching vessel. It was further observed that a grind fineness
of approximately 98 percent passing about 12 microns was
required to achieve approximately 98 percent copper extraction
at about 1700 C. and about 400 kg/tonne sulfuric acid
addition to the pressure leaching vessel.
25
Grinding Size in Microns % ell Residue
System Pso Poo P05 Pos Extracted wt%Cu
1600 C, 500 kg/tonne H2 SO4
30
1 24.6 34.5 43.4 52.2 93.0 2.99
3 6.4 10.8 18.9 31.3 97.0 0.92
4 5.5 8.5 13.7 23.2 98.3 0.72
2 6.7 10.2 17.0 22.3 98.6 0.67
1700 C
35
5* 7.7 11.7 16.5 23.9 96.9 1.39
6** 6.2 7.8 9.2 12.1 98.1 0.76
*500 kg/tonne H2SO4
**400 kg/tonne H2 SO4
40
EXAMPLE 2
Batch results also indicate that copper extraction is sensitive
to grind fineness. The batch tests were performed to 45
confinn that the products of Netzsch mill processing would
react similarly to the products of Metprotech processing.
The grinding systems indicated in Table 2 correspond to the
grinding systems identified in Example I.
50
TABLE 2
Copper Extraction versus Grind Fineness in Batch Tests
Grinding Size in Microns % ell Residue
55
System Pgo Poo P05 Pos Extracted wt%Cu
1600 C, 500 kg/tonne H2 SO4
5 9.8 13.8 18.9 27.8 97.7 0.954
5 9.9 13.6 18.6 28.1 98.4 0.664 60
4 5.7 9.4 13.9 21.3 99.2 0.327
6 6.2 7.8 9.2 12.1 99.2 0.358
1700 C, 500 kg/tonne H2 SO4
5 9.8 13.8 18.9 27.8 95.1 1.930
6 6.2 7.8 9.2 12.1 99.2 0.343
65
US 7,341,700 B2
17
other changes or modifications are intended to be included
within the scope ofthe present invention, as expressed in the
following claims.
What is claimed is:
1. A method for recovering metal from a metal-bearing
material, comprising the steps of:
(a) providing a feed stream comprising a metal-bearing
material;
(b) subjecting at least a portion of said feed stream to
controlled, super-fine grinding to form an inlet stream,
wherein said controlled, super-fine grinding comprises
reducing the particle size of said feed stream to a P98
of less than about 25 microns;
(c) leaching at least a portion of said inlet stream in the
presence of oxygen at an elevated temperature and
pressure to form a metal-bearing solution;
(d) recovering at least one metal value from said metalbearing
solution.
2. The method of claim 1, wherein said step of providing
a feed stream comprising a metal-bearing material comprises
providing a feed stream comprising a copper sulfide
ore or concentrate.
3. The method of claim 1, wherein said step of providing
a feed stream comprising a metal-bearing material comprises
providing a feed stream comprising chalcopyrite.
4. The method of claim 1, wherein said step of subjecting
said feed stream to controlled, super-fine grinding comprises
reducing the particle size of said feed stream to a P98 of
from about 10 to about 23 microns.
5. The method of claim 1, wherein said step of subjecting
said feed stream to controlled, super-fine grinding comprises
reducing the particle size of said feed stream to a P98 of
from about 13 to about IS microns.
6. The method of claim 1, wherein said step of leaching
said inlet stream comprises leaching said inlet stream at a
temperature of from about 140° C. to about 180° C. and at
a pressure of from about 100 psi to about 750 psi.
7. The method of claim 1, wherein said step of leaching
said inlet stream comprises leaching said inlet stream at a
18
temperature of from about 1500° C. to about 1750° C. and
at a pressure of from about 100 psi to about 750 psi.
8. The meted ofclaim 1, wherein said step ofleaching said
inlet stream further comprises leaching said inlet stream in
the presence of a surfactant.
9. The method of claim 1, wherein said step of leaching
said inlet stream further comprises leaching said inlet stream
in the presence of a surfactant selected from the group
consisting of lignin derivatives, orthophenylene diamine,
10 alkyl sulfonates, and mixtures thereof.
10. The method of claim 1, wherein said step ofleaching
said inlet stream further comprises leaching said inlet stream
in the presence of calcium lignosulfonate.
11. The method of claim 1, wherein said step of leaching
15 said inlet stream further comprises leaching said inlet stream
in the presence of a surfactant in an amount of from about
2 to about 20 kilograms per tonne of concentrate in the inlet
stream.
12. The method of claim 1 further comprising the step of
20 combining at least a portion of said inlet stream with a liquid
prior to step (c) to form an inlet stream having a solid-liquid
ratio.
13. The method of claim 12, wherein said step of combining
at least a portion of said inlet stream with a liquid
25 comprises combining at least a portion of said inlet stream
with at least one of process water, raffinate, pregnant leach
solution, and lean electrolytc.
14. The method of claim 12, wherein said step of combining
at least a portion of said inlet stream with a liquid
30 comprises combining at least a portion of said inlet stream
with a liquid to form an inlet stream having a solid-liquid
ratio of from about 5 percent to about 50 percent.
15. The method of claim 12, wherein said step of combining
at least a portion of said inlet stream with a liquid
35 comprises combining at least a portion of said inlet stream
with a liquid to form an inlet stream having a solid-liquid
ratio of from about 10 percent to about 35 percent.
* * * * *