FI58347C - FOERFARANDE FOER UTVINNING AV KOPPAR ZINK ELLER NICKEL FRAON EN MALM INNEHAOLLANDE DESSA I FORM AV SULFIDMINERAL FOERENAD MED JAERNSULFID - Google Patents

Description

Ia] LH YiH ΐιΐ '^ ΙΙΪ Ι -'-' ^ - Ι rBl ANNOUNCEMENT 1: 07/9 -M (11> UTLÄGGNINe $ SKRIPT OÖ047 C (4S) Patent ayonnetty 1 £ Oi 1081 Patent meddelat ^ (s,) Kv .ik.Wa3 C 22 B 3/00 // O 22 B 15/10 FINLAND — FINLAND (21) p ^^ twetoon * 690/73 06.03.73 'ΓΙ' (23) AlkupMvt — GIWtlMttd ^ 06.03.73 (41) ) Tullut JulklMM - MM * offwtHg Οθ. 09.73

Patent and other registered registrar * (44) NUKtviksIptnon j »lu> uL |

Patent · och ragitterttyralaan AmMcm uttagd och mlJikrifteR puMicurad 30.09.80 (32) (33) (31) fnrd «** y« tosikuut — l ^ lrd pHorkm 07.03.72 10.Oi * .72, 23.06.72, 27.10.72 USA (US) 232U5U 2k23k0, 265709, 3011 * 75 (71) The Anaconda Company, 25, Broadway, New York, NY, USA (US) (72) Martin Clifford Kuhn, Tucson, Arizona, Nathaniel Arbiter, Tucson,

Arizona, USA (7I +) Berggren Oy Ab (51 *) Process for the recovery of copper, zinc or nickel from ore containing these in the form of a sulphide mineral associated with iron sulphide - For the recovery of copper from zinc or nickel in the form of a sulphide mineral av sulfidmineral förenad med järn-sulfid

Common practice in the recovery of copper, zinc and nickel from sulphide ores containing them involves flotation flotation of the ores, production of concentrate containing valuable metal sulphides and disposal of non-valuable sulphides and silicates, aluminates and other non-valuable by-products. These concentrates are treated differently, depending on the precious metals present. Copper and nickel flotation concentrates are usually smelted to produce metallic copper and nickel, and zinc flotation concentrates are roasted to oxide and either reduced with carbon (coke) or soaked with acid and reduced to the metal electrolytically.

Immediately, direct soaking of sulfide concentrates has been proposed to prepare a soaking solution from which the precious metal can be recovered either directly by chemical reduction or electrolytically. According to one such method, described in U.S. Patents 2,576,314, 2,726,934 and 2,822,263, nickel or copper concentrates are soaked in ammonia in the presence of 2,58347 air in an autoclave at high temperature and pressure. The resulting soaking solution containing copper or nickel is treated in an autoclave at high temperature and pressure with hydrogen or carbon monoxide to precipitate nickel or copper as a fine metal. The process of the aforementioned U.S. patents has been successfully used to treat relatively valuable nickel concentrates, but has not been economically satisfactory in the treatment of copper concentrates.

The current Finnish patent applications Nos. 689/73, 1080/73 and 1155/73 describe a new method for soaking copper, nickel and zinc concentrates in an aqueous, ammoniacal solution at relatively low temperatures and near atmospheric pressure. The precious metals in the soaking solution are recovered by ion exchange in the liquid and then by electrolysis, or by some other suitable means.

It has also been proposed to treat copper concentrates with concentrated (almost 100%) sulfuric acid to convert metal sulfides to sulfates, thereby releasing some elemental sulfur. The copper in the concentrates thus treated is recovered as an acidic copper sulphate solution from which it can be separated in any suitable known manner. Although this method is technically possible, it has so far not proved to be economically viable.

Except as described above, in all prior art commercial processes for recovering a precious metal by soaking, roasting of concentrates to oxides (resulting in the evolution of large amounts of sulfur dioxide) and soaking of the resulting roast with sulfuric acid or other acid are used. With the exception of the treatment of zinc concentrates for the electrolysis-recovery of zinc, this method has not been found to have an economic advantage over the conventional sulfide smelting technique.

The present invention therefore relates to a process for recovering copper, zinc or nickel from ore containing these in the form of a sulphide mineral associated with ferrous sulphide, the ore being treated by foaming to obtain a primary sulphide concentrate of the mineral and associated ferrous sulphide and dissolved in ammonium sulphate and ammonium hydroxide At a temperature of p and a pressure not exceeding 0.7 kg / cm in the presence of free oxygen and the resulting rich solution is separated from the remaining dissolved primary concentrate. According to the present invention, dissolution is stopped while the remaining dissolved primary concentrate still contains a substantial amount of undissolved copper, nickel or zinc, the remaining dissolved primary concentrates are then subjected to a second flotation of a secondary primary concentrate containing copper, nickel and nickel or nickel. and to provide a waste containing ferrous sulfide from which copper, nickel and zinc minerals have been removed during dissolution. In this way, relatively insoluble undissolved sulfides and other minerals can be discarded with the secondary fraction of the secondary flotation, and a secondary concentrate of better quality is obtained, which is less, and which is excellently suitable for further processing. This secondary concentrate can, depending on the nature of the original or primary concentrate, be treated by melting, re-soaking or any other desired technique. Regardless of which treatment method is chosen, this secondary concentrate is of better quality for the selected treatment than the primary concentrate, and its volume is so much smaller than that of the primary concentrate that it requires only a relatively small treatment plant.

Undissolved solids removed from the soaking supply before being subjected to said secondary foam flotation must be washed free of most of the ammoniacal soaking solution, e.g. The washed and filtered or thickened solid is transferred to a secondary flotation delivery, and the wash water is preferably combined with the soaking solution for recovery treatment of the dissolved precious metal.

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The solid from the soaking delivery can be re-soaked either before or after enrichment in the secondary flotation delivery. For example, if it is desired to recover substantially all of the soluble precious metal (e.g., copper) from the primary concentrate and then recover substantially insoluble precious metal sulfide (e.g., molybdenite) to the secondary concentrate, it may be advisable to soak the solid stream after separation (e.g. a soaking operation to substantially complete the soaking of soluble sulfides before the insoluble residue of the soaking operations is washed, thickened and subjected to a secondary flotation operation. Alternatively, if it is desired to ensure substantially complete recovery of the soluble precious metal by soaking and re-soaking while minimizing the volume of material subjected to secondary soaking, the solid leaching solids pyrite) can then be re-soaked in a secondary soak delivery to substantially complete the soak of soluble metal sulfides.

The process according to the invention can be used with particular advantage for the treatment of copper sulphide concentrates. These concentrates often contain a large number of particles of relatively worthless sulfides, such as pyrite, whose surface is covered only with a thin layer of copper sulfide mineral. Such a concentrate can be treated according to the invention by subjecting it to an ammoniacal soaking supply to the extent necessary to dissolve such surface coatings (and preferably also copper sulfide minerals contained in the primary concentrate mucus components), but not to the extent necessary to dissolve relatively large amounts of copper concentrate. list. Upon completion of such a soaking operation, the solid waste is washed and thickened and then subjected to a secondary flotation operation to produce a secondary concentrate from which the undesired pyrite and mucus have been substantially removed. This concentrate usually takes up much less space and is of better quality than the primary concentrate. It can be processed by melting or re-soaking as desired.

The process according to the invention can also be advantageously used for the treatment of mixed liquids such as copper-nickel or copper-zinc sulphide concentrates. The copper content of the primary concentrate, especially the part which is in the form of chalcosite and covellite, dissolves somewhat more easily than ordinary zinc-nickel sulphide [minerals. The shark is made into a primary soaking solution that is somewhat enriched in relation to copper compared to nickel or zinc (although it contains both copper and nickel or zinc), and after thickening and washing, the solid residue is enriched in a secondary enrichment of nickel or zinc. and correspondingly depleted relative to copper compared to the primary concentrate. Such enrichment of the primary soaking solution with respect to copper and enrichment of the secondary concentrate with respect to zinc or nickel, although not large, may facilitate the separation of the precious metals of the primary concentrate.

When the soaking solution contains both copper and either nickel or zinc, copper can be separated by treating this solution with liquid ion exchangers and leaving the zinc or nickel in the raffinate from which these are. precious metals can be recovered separately from ammonia and ammonium sulfate.

The accompanying drawing shows flow diagrams of a few alternative embodiments of the process of the invention applied to the use of a primary copper concentrate as a feedstock. Substantially the same set of process steps as shown in these flow diagrams can also be applied to primary nickel and zinc concentrates.

Figure 1 is a flow chart of an embodiment of the invention in which a secondary flotation delivery is performed on a solid residue from a secondary soaking delivery. Fig. 2 is a flow chart of an embodiment of the invention in which a secondary flotation operation is performed on a solid residue from a primary soaking supply and in which the secondary concentrate is re-soaked to recover its sulfide precious metal; flotation delivery to make a better concentrate for smelting.

The flow chart of Figure 1 outlines an embodiment of the invention suitable for treating a primary copper concentrate containing mainly copper sulfide rinerals (e.g., caicosite, chalcopyrite, cosellite, bornite, etc.) and small amounts of precious metals such as silver and gold. and in addition a small amount of 6,583,447 value sulfides other than molybdenite and possibly lead wedge In the process of Figure 1, most of the copper (e.g.

95% or more) is recovered by soaking, and a small portion of the copper, along with other precious metals, is recovered in the secondary flotation concentrate.

The primary concentrate is first slurried in a recycled aqueous ammonium sulfate solution containing a small amount of free ammonia. The ratio of solid to liquid is not critical, but may be 10-20, for example 15 wt.%. The concentration of ammonia or ammonium sulfate in the formation of the slurry is not critical. Ammonium sulfate is generated during the process, and ammonia is added during the soaking delivery, so recycled liquid, regardless of its ammonium sulfate content, can be used to form the slurry.

The slurry is then delivered to the primary soaking step, which is preferably carried out as described in the above-mentioned Finnish patent application No. 689/73. Accordingly, the slurry is passed through a series of separate, closed soaking vessels, each having a pressure of less than 0.7 aty, preferably less than 0.6 aty, and a temperature of 50-80 ° C. The temperature is close to the lower limit of this range in the first vessel of the series and rises a few degrees C in each subsequent vessel. The slurry is vigorously agitated in each vessel while feeding oxygen to each vessel, while also feeding ammonia to at least the first or a few first vessels in the series. The amount of ammonia fed is correlated with the temperature of the slurry so that the sum of the water vapor and the partial pressures of ammonia is at least a few tenths of the atmosphere (e.g.

0 »35-0.6 at) total system pressure - not exceeding 0.7 aty - lower. This pressure difference is filled by the oxygen supplied. At the desired temperatures, the amount of ammonia that can be fed into the system without exceeding said limit corresponds to a pH of the slurry between 9-11.

The soaking reaction proceeds rapidly and exothermically in the first or first few vessels of the series, and maintaining the desired temperature may require cooling the slurry with cooling coils or a cooling jacket surrounding the vessel. In the last or the last few vessels, where the soaking progresses more slowly, the ammonia fed from the previous vessels is usually handled, so that there is no need to add ammonia to the last few vessels in the series. It may also be necessary to introduce some heat into the slurry by means of heating coils or a heating mantle to maintain the desired temperature in these vessels.

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Oxygen (preferably commercially pure oxygen) is added to all vessels in the series and, as mentioned above, the temperature and ammonia content of the slurry in each vessel are adjusted so that the sum of ammonia, water vapor and inert gas partial pressures is at least about 0.35 aty less than total system pressure. in the supernatant atmosphere of the slurry in each vessel, there will be at least this 0.35 aty oxygen pressure.

The sludge in each vessel is vigorously agitated, and the oxygen from the atmosphere above the surface of the sludge in each vessel is vigorously circulated quite deep below the surface. This agitation and intense oxygen recirculation can both be achieved with a suitable, powerful lower aeration propeller mounted on each soaking vessel.

The primary soaking can take 3 to 6 hours, typically 5 hours, to allow the slurry to pass through a series of five soaking vessels. During this time, most of the more soluble copper sulfide minerals (e.g., chalcosite and covellite) dissolve, and also at least some of the less sparingly soluble minerals (e.g., bornite, chalcopyrite, and enargite). Copper enters the solution as copper ammonium sulfate. A significant proportion of the less soluble copper minerals and essentially all of the molybdenite and pyrite remain insoluble. Likewise, any lead mica (lead sulphide) present in the soaking residue remains either as lead mica or one of its transformation products. After the primary leaching operation, most of the copper-containing soaking solution is separated from the solid by decantation in a thickener, and the copper solution is passed to a recovery of its copper content. The thickened solid is reslurried in the recycled ammonium sulfate solution and subjected to a secondary soaking treatment that is performed in much the same way as the primary soaking, except usually in a single closed soaking vessel or possibly a series of two vessels. The temperature is then maintained at 65-80 ° C (up to 90 ° C if nickel or zinc concentrates are treated), and both ammonia and oxygen are introduced into the slurry. Even now, however, the ammonia content is kept so low that the sum of its partial pressure and water vapor partial pressure is at least a few (0.35-0.6 aty) tenths lower than the total system pressure, which in turn does not exceed 0.7 aty and is preferably less than 0 , 6 aty. The time required for secondary soaking is typically about 6 hours. As with primary soaking, sludge is strongly agitated in the secondary soaking in the tank (or tanks), and oxygen is circulated strongly from the atmosphere above the surface of the sludge quite deep below the surface of the sludge.

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At the end of the secondary leaching, most of the copper content of the original feed primary concentrate has dissolved, leaving only substantially insoluble sulfides such as pyrite, molybdenite, lead mica or its insoluble reaction products, and perhaps 1-5% of the original copper content, such as most of the original copper content. Even the pyrrhotite that may have been originally present remains largely insoluble. And, of course, side minerals such as quartz and aluminum and magnesium silicates that may have been included in the primary concentrate remain insoluble.

The slurry leaving the secondary leaching delivery is washed countercurrently by decantation to separate the solid from the soaking solution and to form a thickened washed mass from the insoluble solid. The secondary leach solution is combined with the primary leach solution to recover the precious metal dissolved therein, and the thickened, washed solid is subjected to a secondary flotation treatment.

Solid flotation is performed using conventional flotation equipment, reagents, and techniques to enrich the precious metal sulfide present and reject low-grade sulfides such as pyrite. The thickened pulp from the countercurrent decantation wash is thus diluted to a flotation-appropriate consistency, and if it contains some copper 3 ulide minerals as well as molybdenite (or lead mica, or both), glotting can be performed, for example, using a xanthate collector lead enamel, if any), and a wreckage fraction containing most of the pyrite, pyrrotite, and side minerals. The wreckage fraction is taken to a landfill, and the secondary concentrate is treated to recover its precious metal. If it contains lead mica as a precious metal, it may be subjected to yet another selective flotation treatment to recover the lead mica to a separate concentrate, or such separation may be performed during the secondary flotation delivery.

The volume of the secondary concentrate is generally very small compared to the volume of the primary concentrate, and its handling is correspondingly simpler. If the primary concentrate contains e3im. 30% copper and 0.1% molybdenum as molybdenite, the secondary concentrate typically contains about 10% molybdenum and about 30% copper, which makes the separation of molybdenum from copper by conventional molybdenite notation in which copper sinks relatively easy. After separation of the molybdenite, the remaining secondary concentrate can be sent, the copper melting to the pregnant to recover its copper content.

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The combined metal-containing solution of the primary and secondary leaching treatments is preferably treated with a liquid ion exchanger, preferably such that it takes copper from the former solution. Several such liquid ion exchangers are known and commercially available, and any of them can be used successfully. They are usually dissolved in an organic solvent and diluted with a water-immiscible organic diluent such as fuel oil. The metal-containing, aqueous soaking solution is stirred together with this organic ion exchange solution, usually in several successive extraction steps. The aqueous, substantially copper-free raffinate thus obtained contains ammonium sulfate and residual ammonia, and is used as a recycle liquid to form a slurry from the primary concentrate. A side stream is separated from the recycled raffinate stream to remove excess dissolved sulfate during the soaking operation and to recover the associated ammonia. This can be done, for example, by precipitating the sulphate with lime as gypsum and boiling the ammonia to remove the residual solution thus obtained.

The loaded organic phase from the liquid ion exchange treatment is passed to a discharge process where the copper is separated by means of a concentrated sulfuric acid solution (spent electrolyte from a subsequent electrolyte supply). The decomposed, copper-free organic phase is then recycled back to the liquid ion exchange supply as usual.

A loaded sulfuric acid solution containing copper discharged from the ion exchanger is subjected to conventional electrolysis treatment. In it, copper is electrically deposited from solution, using insoluble anodes, into standard shaped copper to be melted and cast into suitable cathodes or to be marketed as such. The spent electrolyte, whose acid content is increased due to electrolytic cell reactions, is recycled back to the ion exchange agent discharge treatment as described above.

The flow chart of Figure 2 outlines an embodiment of the invention suitable for treating a low quality primary copper concentrate containing only copper sulfide minerals, or perhaps these minerals in conjunction with zinc sulfide. The process of Figure 2 uses a secondary flotation treatment to simplify the recovery of the primary concentrate precious metal by secondary soaking.

A primary feed concentrate containing e.g. .20% copper, much of which is in the form of slime and a surface layer of pyrite particles, and perhaps 1% zinc (sphalerite) or a few% nickel (pentlandite), 10 58347 is slurried as described above, back to recycled ammonia. ammonium sulphate solution to a slurry of about 15% solids.

The resulting slurry is subjected to a primary soaking treatment, which, as described in connection with Figure 1, can be performed in a series comprising about five closed soaking vessels. During the soaking, ammonia and commercially pure oxygen are added to the solution. Ammonia only needs to be added to the first one or two vessels in the series, but oxygen is fed to all of them. Temperatures and other soaking conditions are those described in connection with Figure 1, care being taken to ensure that the temperature and free ammonia concentration in each vessel is such that the sum of the partial pressures of ammonia, water vapor and inert gases is at least about 0.35 at less than pressure , which develops by supplying oxygen to the system so that the partial pressure of oxygen in each vessel is substantially equal to this.

Typically, about 80-90% of the copper in the primary concentrate and much of its zinc and nickel dissolves during the primary leaching, resulting in perhaps about 50% weight loss of the solids. , and the solid is recovered as a thickened slurry. The soaking solution is subjected to a liquid ion exchange treatment to recover the precious metals dissolved therein, and the solid is diluted to a suitable flotation consistency and then subjected to a flotation treatment.

The flotation operation is performed in a conventional manner to recover the precious metal sulfide to the secondary concentrate and to discard the side rock constituents and worthless sulfides of the primary concentrate along with the waste fraction. Using conventional xanthate collectors and standard flotation conditioning, the remaining copper sulfide minerals (mainly relatively sparingly soluble minerals such as enargite and chalcopyrite) and insoluble zinc sulfide minerals are made substantially completely collected in about 25% copper, each secondary concentrate. Pyrite, which is free of surface coatings formed of valuable minerals, and the remaining side rock components from which the copper or zinc sulphides adhered to them have been soaked, are discarded with the flotation fraction.

The volume of the secondary flotation concentrate is reduced to as much as about 10-15 i of the volume of the initial feed concentrate and a third or less of the volume of solids obtained from the primary soaking. This small volume of secondary concentrate, which contains # substantially all of the still undissolved copper plus insoluble zinc, is the ideal feedstock for the small secondary soaking treatment to which it is now passed.

Due to the small volume of the secondary concentrate, the secondary leaching can usually be performed in a single compact leaching vessel. The soaking conditions are essentially the same as described above in connection with Figure 1. The temperature is as high or higher than in the final stages of primary leaching to promote substantially complete dissolution of the copper (and zinc or nickel) sulfides. The soaking time can be economically equal to the total duration of the primary soaking, (5 or 6 hours), taking into account the small volume of the secondary concentrate.

After secondary soaking, the slurry from the soaking vessel is filtered (and, if desired, the filtered solid is washed). Even now, due to the small volume of solids remaining after secondary soaking (typically only about 40% of the secondary concentrate), a small-scale and thus relatively simple filtration delivery is sufficient to separate the solids from the soaking solution. The latter is combined with a primary leach solution for treatment in a liquid ion exchange delivery, and the filtered solid is discarded, or if it contains recoverable materials such as noble metals, it can be sent for further treatment.

The treatment of the metal-containing solution obtained from the primary and secondary soaking by solvent ion exchange and electrolysis is essentially the same as described above in connection with Figure 1. During soaking, the dissolved zinc or nickel remains in the raffinate obtained by liquid ion exchange and can be recovered from the branch stream used to remove excess sulfate and recover ammonia. For example, after precipitation of sulfate as gypsum, zinc can be recovered simultaneously with boiling the residual solution to recover ammonia.

Alternatively, zinc can be precipitated by treating the branch stream with carbon dioxide at a pressure (0.7-3.6 aty) to precipitate complex zinc carbonate, which can be filtered or otherwise separated from the solution. The residue can then be gypsum precipitated with lime and then boiled to recover ammonia.

If nickel is present, it can be recovered by treating the entire raffinate or its branch stream separately in an ion exchange process for removing nickel, after which the nickel-free raffinate is treated to remove excess sulfate and recover 12,58347 ammonia. Nickel can also be recovered from the branch stream after precipitation of excess sulphate as gypsum, while the solution is boiled to recover its ammonia content.

It is not necessary to recover all the zinc or nickel from the raffinate. Although quite a few of these precious metals are recycled back to the soaking supply, no damage occurs. All that is needed is to recover so much zinc or nickel that these metals cannot be excessively enriched in the soaking solution.

The flow chart shown in Figure 3 relates to an embodiment of the invention which is well suited for treating poor quality concentrate to improve its quality for smelting or to increase the amount of concentrate that can be processed to recover copper in combination with an existing smelter without actually increasing smelter production capacity or sulfur dioxide output. In the process of Figure 3, secondary flotation delivery is used to make a high quality copper smelter feedstock, thereby reducing the volume of the feedstock, or to increase the amount of primary concentrate to be treated without increasing the weight of concentrate fed to the smelter. Again, the primary concentrate may contain substantially only copper as the recoverable precious metal, or it may contain recoverable amounts of zinc or nickel.

The primary concentrate feed is returned to the recycled ammonia-ammonium sulfate solution as described above. The resulting slurry is then subjected to a soaking treatment, which is performed essentially as described above in connection with Figure 1. The main difference is that the process of Figure 3 does not seek to achieve as much solubility of copper and other soluble precious metals in the soaking solution as the processes of Figures 1 and 2. For example, it may be possible to recover only half or even less of the copper content of the primary concentrate in the soaking solution. Thus, the number of soaking vessels in a series can be considerably smaller, and even a single soaking vessel may be sufficient in some cases. Except for this small amount of soaking equipment, the soaking operation is performed essentially as described in connection with Figure 1. Soaking temperatures can be maintained close to the lower end of said range (50-80 ° C) if the primary dew contains a significant amount of readily soluble copper minerals.

The soaking time can also be shortened in approximately the same proportion as the number of soaking vessels is reduced. For example, a 1: i.o. time of 1-2 hours is usually sufficient to dissolve up to 50% of the copper contained in the primary nitrogen.

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After the soaking delivery, the slurry is passed to a decantation-washing delivery, where the rich soaking solution is separated from the undissolved solid and this is recovered as a washed, thickened sludge. The solution is further sent to a liquid ion exchange treatment to recover its copper content, and the thickened solid is diluted to flotation consistency and subjected to flotation delivery.

This flotation delivery is generally substantially similar to the delivery with which the primary concentrate was prepared. The collector may be, for example, a xanthate, and the conditioning of the flotation mass may be the same as that used in the preparation of the primary concentrate. Undissolved copper sulphide minerals accumulate in the secondary flotation concentrate (together with insoluble zinc or nickel minerals, if any), and excess pyrite and side rock minerals soaked free of copper and other valuable sulphides in the leaching process.

The weight of the secondary concentrate is significantly less than the weight of the primary concentrate originally fed to the soaking supply.

In fact, this reduction may be disproportionately greater than the proportion of copper minerals dissolved during soaking. Thus, if the primary concentrate contains a significant amount of pyrite, the particles of which are coated with copper sulfides on the surface, most of the pyrite is released from its copper mineral coatings during the soaking delivery and can therefore be discarded with the secondary flotation delivery. Such effects, together with a decrease in the volume of the primary concentrate due to the dissolution of the copper minerals therefrom, result in the weight of the secondary concentrate being much less than the weight of the primary concentrate - a decrease which may disproportionately exceed the decrease due to the solubility of the copper minerals alone.

If the primary concentrate contains a significant amount of mucus, the quality of the secondary concentrate may be significantly improved due to the dissolution of the copper mineral mucus in the soaking treatment.

'The result is that the secondary f lotts iorikast e as well. the amount decreases that can improve the quality of the molten feed. less capacity of the concentrator is required to process the secondary concentrate than would be required to process the primary concentrate, which allows the weight of the primary concentrate to be greatly increased without increasing the capacity of the melter; or alternatively a lower suin ti n! ·: the opacity is sufficient to handle a certain amount of primary concentrate. Greater attention can be paid to recovering as much copper as possible from the copper ore to the primary concentrate, without having to worry that the quality of the primary concentrate deteriorates, as the quality of the concentrate for smelting purposes can be improved by secondary flotation treatment.

The rich soaking solution in the process of Figure 3 is treated in the same manner as described above in connection with Figures 1 and 2. If nickel or zinc is included in the primary concentrate, it can, to the extent that it is dissolved in the soaking solution, be recovered from the solvent ion exchange raffinate, as described in connection with Figure 2.