Process for the recovery of precious metals from a roaster calcine leach residue

A process for separating precious metals from a roaster calcine leach residue from a process wherein copper or zinc sulfides are roasted to produce a copper or zinc calcine; the calcine is leached with an aqueous sulfuric acid leaching solution to produce a copper or zinc-containing leaching solution and a roaster calcine leach residue and the copper or zinc-containing leaching solution is separated from the roaster calcine leach residue wherein the process comprises: PA0 (a) intimately contacting the roaster calcine leach residue with an aqueous sulfuric acid leach solution containing from about 5 to about 200 grams per liter of sulfuric acid to produce a slurry of leach solution and roaster calcine leach residue and to dissolve precious metal from roaster calcine leach residue; PA0 (b) adding copper or zinc sulfide solids to the mixture of leach solution and said roaster calcine each residue; PA0 (c) agitating the copper or zinc sulfide solids in intimate contact with the mixture of leach solution and roaster calcine leach residue under oxidizing conditions for a time from about 5 to about 20 minutes to collect precious metal on the copper or zinc sulfide solids; PA0 (d) separating the leach solution from the roaster calcine leach residue and the copper or zinc sulfide solids; and, PA0 (e) separating the copper or zinc sulfide solids from the roaster calcine leach residue by a froth flotation process.

This invention relates to the recovery of precious metals from leach 
residues from roaster calcines from a process wherein metal sulfides such 
as copper or zinc sulfides are roasted to produce a metal calcine which is 
thereafter leached with an aqueous sulfuric acid solution to recover metal 
from the metal calcine leaving a leach residue. 
In many processes for the production of copper or zinc metal from copper 
sulfide or zinc sulfide ores, roasting is used. Similar processes are used 
for the recovery of these metals from their respective sulfide ores and 
they will be referred to herein as metals unless one or the other is 
identified. In the roasting process, the metal sulfide materials are 
converted to metal oxide or metal sulfate materials which are soluble in 
sulfuric acid leaching solutions. The metal calcine resulting from such 
roasting processes is normally leached with an aqueous sulfuric acid 
solution to produce a metal-containing leach solution which is passed to 
further processing for metal recovery with the residual solids being 
passed to disposal or further treatment for the recovery of residual metal 
values. It has been found in such processes that when precious metals such 
as silver and gold are present in the metal sulfide ores subjected to 
roasting, such precious metals are found with the solids after the 
leaching step. In many instances it is desirable that the precious metal 
values be recovered, and a variety of processes have been used to recover 
such precious metal values. Some such processes include treatment of the 
solid residues with cyanide or brines. Both these processes suffer 
significant disadvantages. Cyanide leaching involves neutralizing the 
leach residue with lime to preclude the formation of cyanide gas and, 
further, the base metals (gangues) remaining in the leach residue may 
consume considerable quantities of cyanide. Similarly, brine leaching 
suffers disadvantages in that such systems are extremely corrosive and it 
is difficult to remove the precious metals from brine solutions. As a 
result, a continuing effort has been directed to the development of a 
method whereby such precious metal values can be effectively and 
economically extracted from roasting calcine leach residues. 
In the preparation of the present application, the following references 
were considered: 
U.S. Pat. No. 3,886,257 issued to Snell on May 17, 1975; 
U.S. Pat. No. 3,902,896 issued to Borbely et al. on Sept. 2, 1975; 
U.S. Pat. No. 3,974,253 issued to Snell on Aug. 10, 1976; 
U.S. Pat. No. 4,070,182 issued to Genik-Sas-Berezowsky et al. on Jan. 24, 
1978; 
U.S. Pat. No. 4,111,688 issued to Ichijo on Sept. 5, 1978; 
U.S. Pat. No. 4,138,248 issued to Narain on Feb. 6, 1979; 
U.S. Pat. No. 4,145,212 issued to Bodson on Mar. 20, 1979; 
U.S. Pat. No. 4,152,143 issued to Kausel et al. on May 1, 1979; 
U.S. Pat. No. 4,177,068 issued to Balakrishnan et al. on Dec. 4, 1979; 
U.S. Pat. No. 4,225,342 issued to Freeman et al. on Sept. 30, 1980; 
U.S. Pat. No. 4,266,972 issued to Redondo-Abad et al. on May 12, 1981; 
U.S. Pat. No. 4,269,622 issued to Kerley, Jr. on May 26, 1981; 
"Special Report," C&EN, Feb. 8, 1982, p. 54; 
"New Oxidative Leaching Process Uses Silver to Enhance Copper Recovery," G. 
J. Snell and M. C. Sze, Engineering Development Center, C-E Lummus, E/MJ, 
October, 1977, pp., 100-105, 167; 
"Electrochemistry in Silver Catalyzed Ferric Sulfate Leaching of 
Chalcopyrite," J. D. Miller, P. J. McDonough, and H. Q. Portillo, 
Department of Metallurgy and Metallurgical Engineering, University of 
Utah, pp. 327-338; 
Ore Deposits, Park et al., 2nd Ed., 1970, pp. 478-484; 
"Nature and Properties of Materials," p. 938; "Table 23-2 Standard 
Oxidation-Reduction Potentials and Equilibrium Constants," Pauling, 
College Chemistry, 2nd Ed., 1957, p. 484. 
These references are hereby incorporated in their entirety by reference. 
The article entitled "New Oxidative Leaching Process Uses Silver to Enhance 
Copper Recovery" and U.S. Pat. No. 3,974,253 both discuss the recovery of 
silver from residues from a leaching process. The silver recovery system 
discussed in these two references relates to a system where substantial 
quantities of silver are present in the residue materials and wherein the 
silver is recovered primarily in the form of an aqueous solution for 
recycle as a catalyst. The recovery of residual quantities of silver from 
the process is disclosed to be by way of conventional precious metal 
recovery processes such as cyanide polishing, or the like. 
According to the present invention, it has been found that silver is 
effectively and economically separated from a roaster calcine leach 
residue from a process wherein metal sulfides such as copper or zinc 
sulfides are roasted to produce a metal calcine; said metal calcine is 
leached with an aqueous sulfuric acid leaching solution to produce a 
metal-containing leaching solution and a roaster calcine leach residue and 
said metal-containing leaching solution is separated from said roaster 
calcine leach residue, said process comprising: 
(a) Intimately contacting said roaster calcine leach residue with an 
aqueous sulfuric acid leach solution containing from about 5 to about 200 
grams per liter of sulfuric acid to produce a mixture of said leach 
solution and said roaster calcine leach residue and dissolve precious 
metal from said roaster calcine residue; 
(b) Adding metal sulfide solids to said mixture of said leach solution and 
said roaster calcine leach residue; 
(c) Agitating said metal sulfide solids in intimate contact with said 
mixture of said leach solution and said roaster calcine leach residue 
under oxidizing conditions for a time from about 5 to about 20 minutes to 
collect precious metal on said metal sulfide solids; 
(d) Separating said leach solution from said roaster calcine leach residue 
and said metal sulfide solids; and 
(e) Separating said metal sulfide solids from said roaster calcine leach 
residue by a froth flotation process.

This invention is useful for the recovery of precious metals from roaster 
calcine leach residues from roaster calcines from processes wherein copper 
or zinc sulfides are roasted to produce a copper or zinc calcine which is 
thereafter leached with aqueous sulfuric acid to recover copper or zinc 
respectively from the copper or zinc calcine. For convenience, the process 
will be described in connection with the FIGURE by reference to copper. 
In the FIGURE, a stream containing copper sulfides such as chalcopyrite are 
charged to a fluidized bed roaster 10 via a line 12. Desirably, the copper 
sulfides charged to fluidized bed 10 are relatively concentrated and, when 
chalcopyrite is the primary copper sulfide, concentrations as high as 80 
wt. percent chalcopyrite or even higher may be presented in the stream of 
solids charged to fluidized bed roaster 10. Fluidized bed roaster 10 is 
operated, as known to those skilled in the art, using a suitable free 
oxygen-containing fluidizing gas, such as air, oxygen enriched air, or the 
like, which is charged to fluidized bed roaster 10 through a line 14. 
Exhaust gases, including sulfur oxides and other oxidation products, are 
recovered from fluidized bed roaster 10 through a line 16 and passed to 
scrubbing or other treatment as required for discharge to the atmosphere 
and the like. As well known to those skilled in the art, sulfur values may 
be recovered from the exhaust gas stream. 
A copper calcine containing copper oxide, copper sulfate, sulfide compounds 
and the like is recovered from fluidized bed roaster 10. The copper 
calcine may also contain quantities of Fe.sub.2 O.sub.3 or Fe.sub.3 
O.sub.4. The copper calcine may also contain minor quantities of 
unoxidized copper sulfide materials. Such fluidized bed roasters may 
operate at a temperature in the neighborhood of about 1150.degree. F. and 
generally from about 20 to about 40 percent excess air is used based upon 
the stoichiometric quantities of air required for the oxidation reactions 
in fluidized bed roaster 10. The copper calcine recovered through line 18 
is passed to a leach vessel 20 where it is contacted, preferably with 
agitation, with a leaching solution which is suitably aqueous sulfuric 
acid containing a suitable concentration of sulfuric acid to leach copper 
from the copper calcine. Desirably, oxygen is bubbled through the solution 
in leach vessel 20 in an amount sufficient to keep iron in its solid form, 
i.e., ferric iron such as magnetite or other refractory iron. Typically, 
the copper calcine is treated in leach vessel 20 as a slurry containing 
from about 20 to about 30 weight percent solids. After a suitable leaching 
period, the mixture of leaching solution and solids is passed through a 
line 26 to a liquid/solids separator 28 where a copper containing leaching 
solution is recovered through a line 30 and passed to solvent extraction, 
or other suitable metal recovery, to remove copper and other metals. The 
acid solution after metal removal is desirably recycled to leaching vessel 
20. The solids separated from the leaching solution in liquid/solids 
separator 28 are passed through a line 32 to a leach vessel 34. It has 
been found in processes such as discussed above, which are considered to 
be well known to those skilled in the art, that precious metal values such 
as silver are found primarily with the solids recovered through line 32. 
Since in many instances, these precious metal values represent a 
substantial quantity of precious metal, it is highly desirable that these 
precious metals be recovered. According to the present invention, it has 
been found that precious metals are readily recovered by contacting the 
roaster calcine leach residue in leach vessel 34 with an aqueous sulfuric 
acid leach solution supplied through a line 36 to produce a mixture of the 
leach solution and the roaster calcine leach residue in vessel 34 with the 
mixture being aerated with a free oxygencontaining gas such as air 
supplied through a line 40. The aqueous sulfuric acid leach solution 
supplied through line 36 typically contains from about 5 to about 200 
grams per liter of sulfuric acid and desirably the mixture is maintained 
at a pH from about 1 to about 2. The mixture in vessel 34 should be 
maintained in an oxidizing condition which may be established and 
maintained by sparging a free oxygen-containing gas such as air, oxygen or 
oxygen-enriched air into the mixture in vessel 34 or by the presence of a 
suitable oxidant such as a source of ferric ions or the like in the 
mixture in vessel 34. Conditions in vessel 34 should be oxidizing with 
respect to sulfur. After establishment of oxidizing conditions in vessel 
34, copper sulfide solids are added through a line 38. Desirably, the 
copper sulfide solids comprise chalcopyrite although other copper sulfides 
such as chalcocite, digenite and covellite may be used. it has been found 
that at these conditions in vessel 34, silver is selectively dissolved 
from the roaster calcine leach residue and deposited on the copper sulfide 
solids. While Applicants do not wish to be bound by any particular theory, 
it appears that the precious metals which are normally dissolved to a 
slight degree under such conditions are selectively collected on the 
copper sulfides as more precious metal is dissolved into the solution as a 
result of the depletion of the precious metal in solution by the selective 
collection of the precious metal on the copper sulfide solids. Thus, the 
precious metal is in the process of being dissolved from the leach residue 
and collected on the copper sulfide solids during the time the copper 
sulfide solids are in vessel 34. By contrast to treatments such as cyanide 
leaching, the present process does not require that the precious metals be 
recovered with the leach solution. Further, such residues may consume 
large amounts of cyanide, especially if metals such as copper and zinc are 
present and the use of cyanide in an acid system poses a substantial risk 
of hydrogen cyanide gas generation. Accordingly, solutions which have a 
much lower precious metal content at process conditions can be used since 
the precious metal is selectively removed more or less continuously from 
the solution onto the copper sulfide solids and dissolved more or less 
continuously from the leach residue. Both silver and gold are recovered 
and other precious metals may also be recovered. The present invention is 
particulary well suited to the recovery of silver, although as indicated 
gold is also normally recovered with the silver. The quantity of copper 
sulfide solids added to vessel 34 is suitably an amount effective for the 
collection of the precious metals onto the solids but limited to produce a 
concentrate suitable for precious metal recovery. Such amounts are from 
about 20 to about 50 wt. percent of the leach residue in line 32. While it 
is desirable that the copper solids be maintained in contact with the 
mixture in vessel 34 for an effective time, i.e., from about 5 to about 20 
minutes, it is undesirable that the copper sulfide solids remain in 
contact with the mixture in vessel 34 for longer periods of time. Since 
the selective collection of the precious metals on the copper sulfide 
solids requires the partial dissolution of the copper sulfide solids, it 
is undesirable that the copper sulfide solids be left in vessel 34 for 
unduly long time periods because the copper sulfide solids, especially the 
smaller particles, may dissolve completely with the resulting release of 
precious metals into solution to the detriment of the recovery of the 
precious metals. Typical reaction temperatures in leach vessel 34 are from 
about 80.degree. to about 90.degree. C. The solids comprising the mixture 
of leach residue solids and copper sulfide solids are desirably present in 
vessel 34 as a slurry containing from about 10 to about 60 wt. percent 
solids and preferably from about 25 to about 50 wt. percent solids. 
Desirably, leach vessel 34 is agitated as is leach vessel 20. The mixture 
of leach solution, roaster calcine leach residue and copper sulfide solids 
is removed from leach vessel 34 via a line 42 and passed to a 
liquid/solids separator where the leach solution is recovered and 
optionally passed through a line 52 to form a portion of the aqueous acid 
solution required in leach vessel 20. Desirably, the sulfuric acid 
solution used in leach vessel 34 comprises make-up sulfuric acid which is 
relatively pure by contrast to the sulfuric acid recovered through line 52 
which will normally contain minor quantities of copper. Clearly, such 
quantities of copper will be recovered via line 30 and passed to further 
processing. The solids separated in separator 44 are passed through a line 
46 to a flotation vessel 48. The pH of the solids passed through line 46 
is desirably adjusted prior to the flotation process as necessary. The pH 
can be increased by the addition of lime to the solids and the solids may 
be further ground. The solids are desirably diluted to a suitable 
concentration for the conduct of a flotation process. The neutralization 
or pH adjustment step may take the form of a rinse step followed by a 
neutralization step or the solids may be simply treated with lime or the 
like to increase the pH to a desired level. Conventional froth flotation 
processes may be used in flotation vessel 48 to separate the copper 
sulfide solids from the remaining residues. The remaining residue is 
discarded through line 50. The copper sulfide materials recovered from 
flotation vessel 48 through line 52 contain a major portion, and in some 
instances in excess of 90%, of the silver and gold contained in the copper 
sulfides charged to fluidized bed roaster 10 through line 12. These 
concentrated copper sulfides are passed to further processing for the 
recovery of copper and precious metal values as known to those skilled in 
the art. A further benefit accomplished by the process of the present 
invention lies in the recovery of additional copper sulfide solids through 
line 52. As noted previously, frequently quantities of the copper sulfide 
materials charged to fluidized bed roaster 10 are unreacted in the copper 
calcine. The copper values in such solids are not readily recovered in 
leach vessel 20. As a result, these copper values remain with the solids 
recovered through line 32 and are carried through the process to recovery 
in line 52. 
While the invention has been described by reference to copper, it should be 
understood that the present invention may also be used to recover precious 
metals from leach residues from zinc roasting and leaching processes of 
the same type. The roasting and leaching process conditions may vary 
slightly as known to those skilled in the art. Such variations need not be 
discussed in detail. In vessel 34 the quantity of zinc sulfide solids, 
such as sphalerite and the like, is desirably from about 20 to about 50 
wt. percent based on the weight of the leach residue in line 32. In 
general, the process of the present invention is substantially the same 
for the recovery of precious metals from copper or zinc calcine leach 
residues except as noted above. 
Having thus described the invention by reference to certain of its 
preferred embodiments, it is respectfully pointed out that the embodiments 
described are illustrative rather than limiting in nature and that many 
variations and modifications are possible within the scope of the present 
invention. Many such variations and modifications may be considered 
obvious and desirable by those skilled in the art based upon a review of 
the foregoing description of preferred embodiments.