PROCESS FOR THE SELECTIVE SEPARATION OF BASE METAL SULFIDES AND OXIDES CONTAINED IN AN ORE

Abstract: Process for the separation of ore components by flotation comprising: grinding ore to form pulp, mixing said pulp with sulfide ions and cyanide ions, adjusting the concentration of said sulfide ions to a level at least sufficient to cause depression of base metal mixed sulfides but insufficient to cause substantial activation of pyrites, and adjusting the concentration of said cyanide ions to a level at least sufficient to cause auxiliary depression of the mineral components of said ore which are required to be depressed in said flotation, but insufficient to cause overdepression of said mineral components; said sulfide ions and cyanide ions having been introduced to said pulp at predetermined times and in a predetermined sequence.

Description: PROCESS FOR THE SELECTIVE SEPARATION OF BASE METAL SULFIDES AND OXIDES CONTAINED IN AN ORE FIELD OF THE INVENTION The present invention relates to a process for ore beneficiation by flotation. More particularly, the present invention relates to the direct, i.e., straight, depression and selective flotation (hereinafter also referred to as "sequential flotation") of mixtures of base metal sulfides and/or partially oxidized sulfides (such mixtures being hereinafter referred to as "mixed sulfidςs") in the absence of pH modifiers, such as alkali and acids, which permits normal or better grades and recoveries to be obtained, without incurring the cost of base and acid additives. The applicability of the process of the present invention is not limited to base metal ore beneficiation, but extends also to treatment of other ores, including non-metallic ores and rocks such as coal, which contain base metal mixed sulfides as minor components. BACKGROUND OF THE INVENTION Most of the economically significant base metal ore deposits worldwide contain mixed sulfides. The con- ventional methods for beneficiation of such ores involve, initially, bulk flotation of metal sulfides and/or subse¬ quent selective flotation of each metal sulfide, depending on individual ore characteristics. Oxidized sulfides are. normally recovered separately from non-oxidized sulfides ("consecutive flotation"), since they are not readily floatable except after pretreatment with sulfidizers, to render their surfaces hydrophobic. After such pretreat¬ ment, the oxidized sulfides may also be recovered by flotation. ^^gUREA OMPI . WIPO Conventional selective flotation of mineral sul¬ fide particles requires grinding of the ore to liberation size, formation of an ore pulp, addition of appropriate depressors, activators, collectors and frothing agents and subsequent flotation in multiple stages. Pyrites are some of the most common constituents of base metal ores. Their presence in flotation is unde¬ sirable because they are generally difficult to depress and normally require a relatively highly alkaline medium. Consequently, a great number of industrial scale flotation separations are performed at an alkaline pH obtained by addition of pH modifiers to the pulp, such as lime, soda ash etc. (hereinafter referred to as "alkaline flotation"). Unfortunately, alkaline flotation results in consumption of substantial quantities of such modifiers, and often in consumption of corresponding amounts of pH neutralizers downstream. In addition, high alkalinity often causes overdepression of other valuable components and decreases the efficiency and selectivity of the separation, requiring larger amounts of activators and collectors, and resulting in increased processing costs. As a result of the widespread use of highly alka¬ line flotation media, the flotation behavior of sulfides in such media has been the subject of extensive study which has generated voluminous literature directed to both the theoretical and practical aspects of such flotation. For an overview of the research published on this topic, see Leja, J. (1982), Surface Chemistry of Froth Flotation, pp. 642-659, Plenum Press, New York; and Staff (1982), Flota- tion Review, Mining Engr., Vol. 34, Nos. 3, 4, pp. 275-279, 377-381. However, comparatively little investigation has been devoted to sulfide flotation in the absence of pH modifiers, i.e., at a natural (unmodified) pH determined mainly by the particular ore composition and the quality of the water supply available. Soluble cyanides (such as sodium and potassium) and soluble sulfides such as sodium sulfide, hydrogen sulfide, polysulfides, etc., are commonly used in alkaline ■ - ΕE OMPI flotation as follows: cyanides are used as complexing and depressing agents; soluble sulfides are used (a) as sulfi- dizers for oxides and oxidized sulfides (in "consecutive" flotation of oxides); (b) as sulfide depressants (after bulk flotation and/or prior to selective flotation); and (c) as collector desorbents subsequent to the collection of a floated fraction. If Na 2 S is used, the quantity required for all of the above uses is of the order of 1,000 g/ton of ore or more. Dilute solutions of sodium sulfide (i.e., of the order of 0.1 M) have been used historically by investiga¬ tors to pretreat mineral surfaces preparatory to microflo- tation studies, in order to displace elemental sulfur and other surface oxidation products from sulfide minerals and thereby carefully control experimental conditions, as is necessary in basic research. Such surfaces are thoroughly washed, however, prior to actually carrying out the microflotation tests. One such basic research study was conducted by Y. Nakahiro: Effect of Sodium Sulfide on the Prevention of Copper Activation for Sphalerite, Mem. Fac. Engr. Kyoto Univ., Part 4, Oct. 1978; pp. 241-257. It involved only the investigation of the effect of sodium sulfide and/or sodium cyanide specifically on the copper activation of sphalerite. The sample tested involved extremely pure copper/zinc sulfide from high grade samples further treated to eliminate quartz, galena, pyrite and other impurities. The results indicated that, in that carefully controlled sample and system, small amounts of Na 2 S had a depressant effect on sphalerite, which was enhanced by the copper ion complexing action of NaCN. However, this effect was pH dependent, the author recommending separation of copper from zinc at an alkaline pH above 8.1. Thus, Nakahiro's study was of limited scope and applicability and its results spoke in favor of pH modifi¬ cation to improve selective flotation. U.S. Patent No. 1,469,042 to Hellstrand, issued on September 25, 1923, is directed to a process of bulk (not selective) flotation of a lead-iron (or lead-iron- copper) concentrate using 1-7 lbs of Na~S per ton of mill feed during the wet-grinding stage to accelerate flo¬ tation of (i.e., activate, not depress) the constituents of said concentrate and inhibit that of zinc. There¬ fore, this is not a process of true selective flotation, which involves flotation of one metalliferous constituent at a time and removal thereof before flotation of another metalliferous constituent. In addition, amounts of Na_S used are much higher than in the process of the present invention, and Hellstrand's process is not applied to oxidized sulfides (non-simultaneous, i.e., sequential flotation), the term "flotation of mixed sulfides", as used in this patent, meaning simply flotation of sulfides of several metals, i.e., what is today known in the industry as a bulk concentrate. U.S. Patent No. 1,916,196 to Ayer, issued on July 4, 1933, is directed to a process for simultaneous flotation of mixed copper sulfides (sulfides, oxidized sulfides, and carbonates) using soluble sulfides, such as a^S, as conditioning additives together with other sulfidizing agents at a carefully controlled pH range between 4.8 and 6.5, the objectives being enhancement of sulfidization, precipitation of copper ions from solution and recovery thereof as sulfides, and bulk flotation of all metalliferous mineral particles. A method was sought which would decrease the cost and/or increase the efficiency of selective base metal ore flotation, particularly one which avoids the need for making a large capital expenditure, such as building of new facilities or extensive modification of existing ones. Accordingly, a method was sought which would decrease the number of flotation stages, reduce reagent consumption, and increase flotation selectivity. OBJECTS OF THE INVENTION One object of the present invention is to pro¬ vide a process for ore enrichment by flotation conducted OMPI at an unmodified pH, thereby making it possible to elimi¬ nate the use of pH modifiers such as lime and acids. Another object of the present invention is to provide a process for the depression and selective sequen- tial flotation of base metal mixed sulfides conducted at natural (i.e., unmodified) pH values. Another object of the present invention is to provide a process for the efficient recovery of the mixed sulfides of the individual metals at reduced costs of processing, reagents and equipment, without sacrificing process selectivity or product grades and recoveries. A further object of the present invention is to provide a process for the recovery of base metal mixed sulfides by selective sequential flotation conducted in the absence of pH modifiers (alkaline or acid) but using otherwise conventional types of reagents (collectors, frothers, depressants, activators, etc.) and existing plant facilities and equipment. These and other objects of the present invention will be apparent to one skilled in the art in light of the following description, accompanying drawing, and appended claims. SUMMARY OF THE INVENTION The present invention comprises a process for the separation of ore components by flotation comprising: grinding ore to form pulp, mixing said pulp with sulfide ions and cyanide ions, adjusting the concentration of said sulfide ions to a level at least sufficient to cause depression of base metal mixed sulfides but insufficient to cause substantial activation of pyrites, and adjusting the concentration of said cyanide ions to a level at least sufficient to cause auxiliary depression of the mineral components of said ore which are required to be depressed in said flotation, but insufficient to cause overdepression of said mineral components; said sulfide ions and cyanide ions having been introduced into the pulp at predetermined times and in a predetermined sequence. 1 DETAILED DESCRIPTION OF THE INVENTION The present invention is described in detail in connection with the preferred embodiments and particularly in connection with Fig. 1, which is a schematic flowsheet 5 of a base metal mixed sulfide flotation process, and Figs. 2 and 3, which are schematic flowsheets of Mo-Cu sulfide flo¬ tation processes. A complex base metal ore, comprising mixed sul¬ fides, gangue materials, etc., is subjected to conventional -I Q coarse-size reduction (crushing) and, subsequently, to fine- size reduction (wet-grinding) to reduce the particles of the valuable metalliferous components to liberation size. This wet-grinding stage may be conducted in one or more stages using conventional equipment (rod, ball or autogeneous mills) - j g to create "ore pulp". Preflotation conditioning according to the present invention may begin as early as the wet-grinding stage, or even slightly before wet-grinding, and may end as late as immediately prior to the first flotation step in the sequence. In Figure 1 , preflotation conditioning can encom¬ 0 pass stages I and II, and more specifically it may include the portion of the Fig. 1 diagram from point 1 to point 2. One aspect of such preflotation conditioning in¬ volves addition of a small amount of sulfide ions (cleanser/ primary depressor) to the ore, preferably during the wet- grinding stage, to achieve better mixing and surface contact 5 and most preferably before any other additives are introduced in the pulp. However, addition of a water-insoluble collec¬ tor at this wet-grinding stage, which is often desirable to reduce overall collector consumption, does not normally affect the sulfide ion action. 0 Another aspect of preflotation conditioning accord¬ ing to the present invention involves addition of a small amount of cyanide ion in the pulp during preflotation condi¬ tioning. Cyanide ion is preferably added after wet-grinding. It is to be noted generally in this discussion 5 that the particular amounts of sulfide and cyanide used in accordance with the present process, as well as the timing and sequence of their introduction, are determined sepa¬ rately for each case because they depend on the particular characteristics (metal and non-metal constituents) of each ore and the quality (mineral content and temperature) of the water employed in its treatment. Thus, for most base metal sulfide ores, sulfide ion is preferably added first, during wet-grinding, followed by cyanide during the remainder of preflotation conditioning. However, cyanide may also be added either simultaneously with the sulfide, or immediately after the end of wet-grinding, or even before addition of the sulfide or in multiple stages. Accordingly, prior to large scale application of the present process to a particular ore, laboratory batch flotation studies should be conducted. These tests may be carried out by first trying concentrations of sulfide and cyanide based on concentrations that previous experience has shown to be suitable for similar ores, or, if there is no previous experience, based on the general ranges dis¬ closed herein, varying said concentrations, until a trend is established, and following that trend until a concentra¬ tion or a concentration range is found that produces optimum results, such as flotation selectivity, increased recovery etc. Suitable sulfide or cyanide ion sources include any reagent which releases sulfide or cyanide ion into an aqueous solution, directly or pursuant to a reaction in the process conditions. Sodium sulfide and sodium hydro- sulfide are preferred, with a 2 S being most preferred. Of the soluble cyanides, sodium cyanide and potassium cyanide are preferred with NaCN being most preferred. Addition of sulfide ion, which in Figure 1 takes place during STAGE I, effects a cleansing of the ore particles during grinding which serves to selectively deoxidize mixed sulfide particle surfaces and to prevent oxidation of freshly exposed surfaces. This facilitates floatability of the mixed sulfide particles during later stages. The ability of sulfide ion to act as a primary depressant of sulfides, which is the second reason for its addition, is also enhanced by its addition during this preflotation conditioning treatment. OMPI WIPO Cyanide ion action is considered to complement sulfide ion action and to enhance selective auxiliary depression of the desired minerals. In addition, cyanide ion serves to complex metal ions in solution. As stated above, the amount of sulfide ion required to obtain both a surface cleansing effect and a primary mixed sulfide depression effect in base metal sulfides depends mostly on ore charateristics (as well as on water quality). If sodium sulfide is used as the source of sulfide ion, the amount required usually ranges between about 20 and 200 g/ton for most base metal sulfide ores. Too small an amount of sulfide ion will be ineffec¬ tive as a depressant (a smaller amount would be also ineffective as a surface cleanser) and too large an amount will cause premature activation of certain sulfides, notably pyrite and in some cases copper, which is generally undesirable in selective flotation processes, in addition to being economically unattractive. As previously men¬ tioned the sulfide ion quantity for each particular application is subject to optimization, which may be indicated by batch flotation testing. It is most preferable to operate a process using the minimum amount of sulfide ion that will produce the desired results (usually between about 20 and 50 g/ton if Na 2 S is used), as use of larger amounts is not only unnecessary (and costly) but it may actually be deleterious to the effectiveness of the present process, by causing a reversal of the depression effect, as discussed above. From the wet grinding stage, the liberated pulp fraction is subjected to a conditioning stage comprising the second portion of preflotation conditioning and labelled "STAGE II" in Fig. 1. Therein, the pulp is conditioned with cyanide ion, preferably NaCN, which serves as an auxiliary depressor, mainly for pyrite, without overdepressing other minerals. Sodium cyanide consumption requirements usually range between about 20 and 200 g/ton, again depending on ore characteristics and process conditions, as was the case with the Na 2 s consumption requirements. Preferred NaCN consumption ranges from about 25 to 100 g/ton. For extremely slimy ore, the addition of a dispersing agent such as sodium silicate with the cyanide can be beneficial. Pulp from STAGE II is further conditioned with collectors and frothers in accordance with usual practice for modern selective flotation in STAGE III. Selective flotation of base metal mixed sulfides in accordance with the present invention begins directly without a bulk flotation step. Thus, the present process is a process of truly sequential (selective) flotation. Depending on ore compo¬ sition, such selective flotation is conducted in the following order from left to right: Pb-[Ag] : Cu : Zn : Fe in accordance with the scheme of Fig. 1 or: Mo : Cu : Fe in accordance with the schemes of Figs. 2 and 3: each metalliferous constituent is activated with an appropriate quantity of a specific activator and/or floated after ad¬ dition of an appropriate quantity of a specific collector (and frother). The process is repeated until a non-float is obtained which, if desired, can be essentially sulfide- free. It is found that by use of the present invention, lower amounts of activators, collectors and frothers are necessary for flotation, as compared to flotation processes of the prior art. If zinc is present in the complex mixed sulfide ore, it must be activated with, e.g., CuSO. prior to flotation. If both zinc and copper are present, the zinc sulfide is likely to be coated with copper ions which would ordinarily render differential flotation of copper from zinc difficult. However, the process of the present invention also solves this problem by complexing and/or desorbing the copper ions from the zinc sulfide surface. The depression effect of the sulfide/cyanide ion combination is transient. Once a metal constituent has been floated and removed, the next one in the sequence can / OMPI be floated easily using the conventional flotation scheme. The transience of sulfide ion action makes it desirable to control the timing of the sulfide ion introduction as well as that of the cyanide ion. However, as mentioned before, this can only be accomplished on a case-by-case basis. The present invention permits one or more of the following major benefits to be obtained. 1 ) Reduction of reagent costs due to pH modifier elimination, use of a relatively small amount of sulfide and cyanide ions, and/or use of reduced amounts of collec¬ tors, activators and frothers. 2) Improvement in flotation selectivity . This permits reduction of operating and equipment costs and further reduction of reagent costs. 3) Improvement in recovery over conventional methods. 4) Improvement in concentrate grades obtained. * 5) Reduction in residence times for conditioning and flotation. 6) Reduction or elimination of deleterious effects which high consumption of flotation reagents may have on further separation of other minerals (e.g. the presence of Ca ions is known to affect the subsequent flotation of cassiterite) . In addition, the present invention makes it possible to increase recovery of extremely fine mixed sulfide particles (slimes) which are normally lost in conventional processes. The present invention, makes it unnecessary and in fact undesirable to add a pH modifier, such as lime, to the pulp. Lime has been customarily added in the wet- grinding stage of base metal ores. It has been found that addition of lime (increasing the pH) actually inhibits optimization of certain steps such as zinc activation. Without the lime, it is possible to operate at the pH range at which copper ion adsorption on zinc mineral particles is at a maximum. These optimization considerations aside, it is generally possible to operate the present process and to obtain its major cost-saving benefits at a pH naturally ranging from about 5.5 to about 8.5. The unmodified pH of a flotation system may vary because of ore composition and local water quality. The important factor here is that pH need not be closely controlled or even monitored, and thus the present process is relatively pH-independent. The present process is applicable to a variety of base metal mixed sulfide ores including, but not limited to, zinc, lead-zinc, lead-zinc-silver, lead-zinc-copper, copper-zinc, and copper-molybdenum. It is also applicable to other ores or rocks such as coal which contain sulfides as minor constituents. In particular, the present process makes it possible to separate molybdenum from copper by straight selective flotation of a molybdenite-rich Cu-Mo concentrate and subsequent flotation of the remaining copper minerals. As is well-known, Cu-Mo combined concentrate is normally floated in one step in primary flotation and is subsequently sent to another plant for further separation. The standard procedure for such separation is to depress the copper and float the molybdenum. Commonly used de¬ pressants in this secondary flotation circuit include any one or combinations of: NaHS, Fe(CN) 2 , NaCN, Nokes' reagent (-P 2 S 5 ^ n N a° H ) an d arsenic Nokes ( s-O, in Na 2 S) . Consumptions of such depressants are generally very high, ranging from about 10 to about 50 kg/ton. Unfortunately, the agents which depress copper also tend to depress molybdenum. Consequently, the Cu-Mo separation requires a relatively large number of stages. Another difficulty stems from the fact that the Cu-Mo concentrate, which becomes the feed in the Cu-Mo separation circuit, is contaminated with collector from the primary circuit, which inhibits later copper depression and necessitates use of large amounts of copper depressants. In order to increase depressant effectiveness and curb secondary circuit reagent consumption, a number of stratagems have been employed to change the surface energy of the copper mineral particles by removing or rendering innocuous the collector coating using procedures such as steaming, roasting or aging of the pulp. It has further been found that use of the present invention in connection with molybdenum containing ores not only affords the benefits enumerated above, and more or less common to all primary flotation circuits, but also makes possible flotation of a Cu-Mo concentrate which is (a) much lower in copper content, and (b) free of a copper collector. This means that the secondary separation (a) will be simplified requiring a smaller number of cleaner stages (and/or resulting in better concentrate grades and recoveries), and (b) will become substantially more cost effective requiring lower (both overall and per-stage) reagent amounts and smaller scale processing equipment. Thus, when the present invention is used, in the pretreatment of a Cii-Mo containing ore, a choice of procedures is available at the copper flotation step as outlined in Figures 2 and 3: (1 ) A collector may be added subsequent to use of the present invention, at point 21 in Figure 2, to •obtain flotation of a substantial volume of a Cu-Mo concentrate following the universal current practice. This procedure will afford one or more of the benefits previously enumerated above. The thus obtained Cu-Mo concentrate will contain most of the Mo and a substantial portion of the Cu (as much as about 90% of the copper and wol contained in the feed), but it will have a very low Mo grade. The concentrate will have to be sent to a conventional Cu-Mo separation plant for further separation. (2) Alternatively, with specific reference to Fig. 3, the copper collector may be omitted, in which case a much lower volume of a Cu-Mo concentrate will be natural¬ ly floated, requirinσ the simple addition of a frother, 31, which may be added substantially simultaneously with the cyanide ion, or at any time thereafter prior to OMPI flotation, 32. The recovery of moly may be the same as in (1), but even if it is lower, the molybdenum grade of the concentrate will be substantially higher (as much as ten times that of (1), above) and the concentrate volume will remain substantially lower than in (1). This concen¬ trate will also need to be sent to a separate plant for further processing but such further processing may be un¬ dertaken directly (without collector removal) and will re¬ quire fewer stages, smaller scale processing equipment, and substantially smaller amounts of Cu-Mo separation depress¬ ants. With continuing reference to Fig. 3, Non-float, 33, which still contains recoverable amounts of Mo is conditioned in accordance with conventional practice with a collector. A further Mo-Cu concentrate, 34, is thus obtained which may be subjected to conventional separation processes. ϊhus, use of the present invention in connection with concentration of a Cu-Mo containing ore, affords added advantages, over processes of the prior art (insofar as the first Mo-Cu concentrate, 32, is concerned). It has been determined in practice that the sul¬ fide ion amount required for primary flotation of a typical Cu-Mo ore in accordance with the present invention varies with the particular ore composition and water quality. If Na 2 S is used as the source of the sulfide ions, the amount required usually ranges between about 5 and 30g/ton, i.e., it is much lower than that generally required for concentration of other base metal mixed sulfide ores such as Pb-Zn. Moreover, the same sulfide ion is used to reactivate the copper minerals after the Mo float is removed. The consumption of cyanide ion is generally the same as in pretreatment of other sulfide ores. Regarding the sequence and timing of sulfide/ cyanide introduction, in Cu-Mo containing ores, it is possible to state generally that introduction of the cyanide preferably follows that of sulfide and involves a distinct step in the process. Another economically advantageous application of the present invention is in coal flotation. Coal is often contaminated by sulfides which are sometimes removed by floating the coal in a conventional process using alkaline flotation. The present invention makes it possible to eliminate alkaline flotation, depress the mixed sulfides, and float coal inexpensively and with high selectivity. EXAMPLES The present invention and its technical and economic advantages are further illustrated by the follow¬ ing examples. These examples in no way limit the scope of the present invention. The laboratory tests were conducted using 1-10 kg portions of different ore samples and standard laboratory facilities, and following the general procedures described above (STAGES I-III) . Tests were run at various locations to test performance of the present invention for a variety of ores and under a variety of local conditions, such as water quality. The pH values obtained during different stages have been recorded. There has been no attempt to change or modify the pH. The values obtained are solely due to ore composition and water characteristics, the effects of any reagents or additives being minimal, due to the low quantities thereof. The pH values obtained in the tests described below ranged between 5.5 and 8.5, showing that (contrary to the generally accepted thinking and practice) operabili- ty of the process is not particularly sensitive to pH changes over a substantial range. Results were generally more favorable at the lower pH end of the above range. The following examples demonstrate that by use of the present invention low cost flotation recovery of mixed sulfide ores, as well as unoxidized sulfide ores, to yield commercial concentrates is possible. The data reproduced below are representative of the tests conducted, including initial tests, and have not been screened. Con- sequently, some of the final values which are less satis¬ factory than others are due to parameters independent of the invention, such as lack of experience of the operators. ORE A - Sample from high-grade oxidized dumps containing about 35% pyrite, 25% argentiferous galena, 15% sphalerite and 25% quartzite ganque. (Villazon-Mojo Area, Potosi, Bolivia). The following tests represent research performed to obtain separate lead-silver and zinc concentrates, from several oxidized dumps considered as potential feed for a custom mill project. The excessive oxidation of the dumps material and the large amount of lime which would have been required to depress pyrite, made the ore difficult to treat and its exploitation non-profitable, prior to use of the present invention. The testing results with comminution to 80% passing 150 mesh are summarized in Table 1, below and show high flotation selectivity and recoveries for all co pon- ents (Zn contained in the Pb-Ag rougher concentrate is recycled into the flotation circuit): OMPI TABLE 1 TEST REAGENTS (g/ton) No. Na 2 S NaCN Na 2 Si0 3 A _ 2 42- CuSO, Z-11-/ A-77 1/ E-ϋ 1 200 150 100 75 300 50 25 6.5 2 150 200 100 75 300 50 25 6.3 3 100 250 100 75 300 50 25 6.2 w No. PRODUCTS %wτ L E A D S I L V E R Z I N C c % Dist.% Oz/t Dist.% % Dist.% 09 0) 1 Pb-Ag Ro CONC. 30.69 60.20 95.65 47.77 94.60 11.27 41.31 H Zn Ro CONC. 15.32 2.45 1.95 3.10 3.08 31.23 57.14 C Non-Float 53.99 0,86 2.40 0.67 2.32 0.24 1.55 H Fl Feed 100.00 19.32 100.00 15.50 100.00 8.37 100.00 ω 2 Pb-Ag Ro CONC. 33.12 62.21 95.76 46.93 94.67 8.02 29.94 Zn Ro CONC. 23.88 2.60 2.88 3.07 4.46 25.55 68.75 5 Non-Float 43.00 0.68 1.36 0.33 0.87 0.27 1.31 Feed 100.00 21.52 100.00 16.42 100.00 8.87 100.00 3 Pb-Ag Ro CONC. 31.44 65.84 94.71 55.37 94.39 5.72 20.34 Zn Ro CONC. 21.56 3.34 3.29 3.93 4.59 32.23 78.54 Non-Float 47.00 0.93 2,00 0.40 1.02 0.21 1.12 Feed 100.00 21.86 100.00 18.44 100.00 8.85 100.00 Dithiophosphate sold by American Cyanamid Corp. 1/ Xanthate sold by Dow Chemical Corp. (iso-propyl) erican Cyiana id Corp. Distribution. Note: The above data fulfill project requirements which did not call for complete separation of lead from zinc. Therefore, the above results are not the product of an optimized separation. ORE B - Sample from oxidized dumps, containing about 30% pyrites, 8% sphalerite-marmatite, 1% cassiterite, 0.5% copper sulfides and siliceous gangue (Milluni Mine, La Paz, Bolivia). The following tests were performed to separate zinc and pyrite to obtain a sulfide-free non-float fraction for subsequent tin (Sn0 2 ) flotation separation. Selective wet grinding in the presence of a 2 S was performed to obtain about 80% passing 150 mesh (105^), i.e., acceptable tin (Sn0 2 ) liberation. Reagent consumption and results appear in Table 2, below. The results show substantial separation of ore components, which had not been possible by use of conven¬ tional processes. OMPI TABLE 2 27 150 200 200 150 150 10 10 30 200 250 250 200 75 10 10 35 100 250 250 100 75 10 - TEST Zn ROUGHER CONC. PYRITE RO. CONC. TIN FLOT. FEED No % T %Zn %DIST T %Sn %DIST %WT Sn %DIST (A 27 32.04 10.96 92. .12 26.47 0.19 9.84 41.49 0.92 74.68 G 30 23.05 13.43 86. .76 32.45 0.15 9.18 44.50 0.91 76.38 09 (Λ 35 13.44 23.53 84, .50 — - - 86.56 0.55 90.82 H ■Ml Test 35 was repeated, using in addition two upgrading (cleaner) stages and a total -4 C of 10 g/ton NaCN. The results were as follows: H PI PRODUCT % T %Sn %Zn DISTRIBUTION 0) Sn Zn X rn ZINC CONC. 8.80 0.35 43.60 5.60 85.94 ZINC MIDDS. 10.40 0.25 2.93 4.73 6.82 ZINC PRIMARY CONC, 19.20 0.30 21.57 10.33 92.76 NON-FLOAT 80.80 0.61 0.40 89.67 7.24 HEADS 100.00 0.55 4.46 100.00 100.00 NOTE - Flotation pH values for all above tests ranged between 6.0 and 5.5, the pH decreasing in the later stages, as expected. The above project became economically more attractive due to the use of the present invention, which resulted in substantial reduction in equipment costs, as well as processing costs. ORE C - Sample from run of mine mixed sulfides containing: 20% sphalerite-marmatite, 30% pyrites and other iron sulfides, 2% boulangerite and jamesonite (lead-silver sulfosalts), and sericitic-quartzitic gangue (Huari-Huari Mine, Potosi, Bolivia). The testing procedure with this ore involved wet grinding in the presence of Na 2 S to 80% passing 150 mesh followed by selective separation of Pb/Ag sulfosalts-zinc concentrates-pyrites (TABLE 4). In subsequent tests, flotation of combined concentrate (sulfosalts and zinc) followed by flotation of pyrite, was effected. (TABLE 5). The reagents employed are summarized in Table 3 below: TABLE 3 TEST R E A G E N T J S (g/to:n) No Na 2 S_ NaCN Z-200 Frother CuSO^ Z-11 I Na 2 Si0 3 2 100 180 50 20 300 50 3 100 120 50 20 150 50 4 150 120 50 20 200 50 5 200 120 50 20 200 50 8 125 150 50 20 300 50. 50 10 100 150 50 20 300 50 50 NOTE - Flotation pH values for all above tests ranged between 6.5 and 5.5. A combined concentrate was obtained in this example because the current plant flowsheet would not permit sulfosalt-zinc selective separation. Thus, the present results in no way reflect on the ability of the present process to effect such selective -separation. However, the ability of the present process to induce substantial recoveries is apparent. OMPI TABLE 4 SELECTIVE SEPARATION PRODUCT OBTAINED TEST SULFOSALTS ZINC ZINC ZINC RO. PYRITE NON- NO. RO. CONC. CONC. MIDDS. CONC. RO. CONC. FLOAT FEED 2 %wτ 3.40 13.97 8.21 22.18 17.49 56.92 100.00 %zn 8.61 45.67 17.50 35.25 0.35 0.30 8.34 cn %DIST 6.54 83.07 5.31 88.38 2.12 2.96 100.00 a: 5,16 55.54 100.00 m 5 % T 15.49 6.97 22.46 16.83 m %Zn 9.42 49.32 8.05 36.51 1.87 0.46 9.26 %DIST 5.25 82.52 6.06 88.58 3.41 2.76 100.00 TABLE 5 COMBINED CONCENTRATES PRODUCT OBTAINED TEST COMBINED COMBINED COMBINED PYRITE NON- NO. CONCENTRATE MIDDS. PRIM. CONC. CONC. FLOAT FEED 8 %WT 16.16 8. .66 24.82 25.30 49.88 100.00 %Zn 49.31 9. .13 35.29 1.64 0.75 9.55 %DIST 83.43 8. .30 91.73 4.35 3.92 100.00 10 %WT 17.50 6. .96 24.46 22.50 53.04 100.00 o 4 %Zn 47.80 6. .44 36.03 0.79 0.45 9.23 § tτ.1 , %DIST 90.64 4. .86 95.50 1.92 2.58 100.00 Based on the results outlined in TABLES 4-5 above, the system has been tested on a commercial scale in a 200 TPD processing plant located at Don Diego, Potosi (Bolivia). The flowsheet of Fig. 1 was used. No special requirements were necessary for startup other than addition of Na 2 S, omission of lime, and minor adjustment of the remaining reagents. The results obtained on this commercial applica¬ tion after two days of continuous testing are shown in Table 6, below: TABLE 6 DAILY MILL REPORT P E R C E N T Z I N C DATE SHIFT HEADS CONC. TAILS % RECOV. 3/26 I 5.69 48.00 0.50 92.17 II 5.33 48.90 0.86 85.37 III 5.48 44.38 1.31 78.41 3/27 I 6.09 47.50 0.65 90.57 II 6.04 47.09 0.65 90.49 III 6.19 49.50 1.11 83.95 NOTE - average pH values ranged between 5.8 and 6.2. A comparison between the present invention and a conventional system in the same plant is set forth in Table 7. The figures for the "conventional lime system" represent the average of January 2 - March 24, 1982 while the figures for the present invention represent the average of the two days' continuous run, described above. This discrepancy in statistical basis should be taken into account when the results in Table 7 are examined. TABLE 7 COMPARISON OF REAGENT SAVINGS (ZINC AND PYRITE SECTIONS) CONVENTIONAL UNMODIFIED pH REAGENT LIME SYSTEM PRESENT INVENTION Price Cost Cost 5 1 $/kg $/τ g $/T CuSO4. 720 0.77 0.554 400 " 0.308 Z-200 19 4.79 0.091 40 0.192 Z-11 100 1.53 0.153 60 0.092 ^^Ε E { OMPI NaCN 26 1.80 0.047 100 0.180 Frother 42 1.38 0.058 42 0.058 Lime 7,500 0.14 1.050 — Na 2 Si0 3 67 0.37 0.025 67 0.025 Na 2 S — 0.80 150 0.120 TOTAL 1.978 0.975 Based on the evaluation of above results, which show substantial cost savings without sacrifice of product grades and recoveries (see Tables 8 and 10 below) the present invention has been in continuous commercial use since May, 1982 at this Potosi plant. Random daily plant data from this commercial application are set forth in Table 8, below. The last entry represents a cumulative average after 21 days' operation. TABLE 8 DAILY MILL REPORT P E R C E N T Z I C DATE SHIFT HEADS CONC. TAILS % RECOV. 5/27 I 6.06 47.56 0.65 90.51 II 5.96 49.96 0.25 96.29 III 5.26 50.46 0.25 95.72 5/28 I 6.11 47.36 0.55 92.07 II 6.46 46.76 0.50 93.26 III 6.46 44.76 0.25 96.67 6/03 I 6.56 48.43 0.57 92.40 II 5.99 50.80 0.41 93.91 III 5.63 48.95 1.14 81 .65 6/21 I-III 7.06 49.23 0.93 88.50 JUNE CUMULATIVE AVERAGE (1-21) 6.42 47 . 1 1 0.72 90. 1 6 The observed variations in reagent consumption were expected as incident to start-up. They were due to factors independent of the present invention, especially the operators' lack of acquaintance with the new proce¬ dures. For this reason, the recent average reagent con¬ sumption, set forth in Table 9 below, is a more meaningful parameter. Consumption of a 2 S shows a reduction of 56% OMPI in Table 9 compared to Table 7. In addition, system optimization reduces consumption of the other reagents. As close monitoring of pH values is no longer necessary in plant operation, pH measurinq equipment and facilities may be eliminated from plants using the present invention. TABLE 9 CURRENT REAGENT DATA - i AVERAGE JUNE, 1982 (ZINC AND PYRITE SECTIONS) COST REAGENT g/ton $/ton CuS0 4 563 0.434 Z-200 44 0.211 Z-11 66 0.101 NaCN 102 0.184 Frother 66 0.091 Na 2 Si0 3 40 0.015 Na 2 S 66 0.053 TOTAL 1.088 Updated data for the above plant based on com¬ mercial operation from June to October 1982 and comparing performance of the circuit utilizing the present process to that of the conventional (lime) circuit ore set forth in Table 10 below: TABLE 10 Lime Circuit - % Zn % Month Tonnes Heads Conct. Tails Recovery Jan. 1982 4456 6.76 50.37 1.19 84.40 Feb. 2494 9.44 49.98 1.27 88.80 Mar. 3427 7.07 47.40 1.17 85.56 Apr. 3723 6.11 48.96 1.43 78.90 May 3127 6.52 47.06 1.39 81.07 Avg. 3445 7.03 48.82 1.29 83.87 No-Lime Circuit Jun. 3035 6.51 47.36 0.77 89.67 Jul. 3137 7.08 45.94 0.77 90.63 Aug. 3694 6.93 47.50 0.68 91.50 Sep. 2957 7.43 48.86 0.76 91.20 Oct. 3609 6.82 49.89 0.77 90.10 Avg. 3286 6.95 47.91 0.75 90.74 ORE D. Sampl( ≥ of run of mine. mixed s ;ulfides containing: 20% sphalerite, 3% galena (6oz Ag per ton), 40% pyrite and siliceous gangue. Liberation size (Zn) is about 80% passing 100 mesh (Porco Mine, Potosi, Bolivia). Differential flotation effects (Pb - Zn) were observed during preliminary testing. However (as in the case of "Ore C", above), such separation was not sought, due to lack of required equipment in the plant. Combined concentrates (Pb + Ag + Zn) were floated from pyrites and gangue, at unmodified pH of 6.5 under the conditions summarized in Table 11 , below and with the results set forth therein. TABLE 11 BATCH FLOTATION TESTS TEST R E A G E N T S (g/ton) NO. PRODUCT % WT % Zn % DIST Na 2 S NaCN CuSO^ 1 Zn Ro. Cone. 21.00 31.63 61.49 100 150 250 Zn Sc. Cone. 36.24 10.95 36.73 Zn Prim. Cone. 56.24 18.87 98.22 Non-float 42.76 0.45 1.78 Heads 100.00 10.61 100.00 I 2 Zn Ro Cone. 26.75 33.34 / 83.40 50 100 250 Zn Sc. Cone. 26.65 5.70 14.20 Zn Prim. Cone. 53.40 19.54; 97.60 I t Non-float 46.60 0.55I 2.40 Cπ I Heads 100.00 10.69 100.00 3 Zn Prim. Cone. 32.37 29.76 86.62 50 125 225 Non-float 67.63 2.20 13.38 Heads 100.00 11.12 100.00 4 Zn Ro Cone. 23.14 27.79 59.28 75 125 250 Zn Sc. Cone. 18.74 18.92 32.68 Zn Prim. Cone. 41.88 23.82 91.96 ! Non-float 58.12 1.50 8.04 Heads 100.00 10.85 100.00 The collector was Z-200 and the frother was "Dowfroth 250", a polyglycol ether (polypropylene ether) sold under this trademark by the Dow Chemical Corporation. Consumption of each was 40 g/ton. Conditioning and flotation times were 5 and 10 minutes per stage, respectively. No upgrading tests were performed. The above results, which show substantial flotation selectivity and recoveries at optimum or near optimum Na-S, NaCN and CuSO, concentrations, formed the basis for a plant testing program at 400 TPD, during 5 days, with the following results: TABLE 12 PLANT TESTING - - CONDITIONS AND RESULTS (Flowsheet as per Fig. 1) TEST NO. 1 2 3 4 5 LIME SYSTEM REAGENT C O N S U M P T I 0 N ( g / t o n ) Na 2 S 50 55 55 85 60 Z-200 50 50 70 70 70 38 NaCN 75 50 70 50 60 3 CuSO. 300 420 270 360 360 672 D-250 45 15 15 15 15 36 Z-11 85 LIME 11 ,086 PRODUCTS (% Zn) HEADS 9.64 9.74 9.84 10.44 12.11 10.39 CONCENTR. 48.99 51.65 53.63 50.16 54.19 53.08 TAILS 2.15 2.10 3.10 2.97 1.00 1.26 RECOVERY (%) 81.26 81.76 72.70 76.05 93.47 91.56 For comparison purposes, the last column shows plant data obtained under the conventional (lime) system during March, 1982 (monthly average). ORE E - An unknown mixed sulfides sample from Mexico was tested at Mountain States Laboratories (Tucson, Arizona) in February, 1982. TABLE 13 TEST CONDITIONS AND RESULTS GRIND: 80% passing 100 mesh — FLOTATION pH = 8.5 (due to ore composition and water condition at testing facility) . W C 03 REAGENTS (g/ton): Na 2 S 100 CuSO.: 200 NaCN 150 Z-200: 40 m -I a 2 SiO 3 100 D-250: 20 c Conditioning and flotation times were 5 min. and 10 min. ro per stage, respectively. ω x oz/ton % D I S T R I B U T I O N 3 PRODUCT % WT % Pb hSL % Zn Pb Zn Pb-Ag Ro. Cone. 7.46 26.60 24.53 6.6 97.97 80.89 14.83 Zn Ro Cone. 15.60 0.14 1.75 18.0 1.08 12.07 84.59 Pyrite Ro. Cone. 2.84 0.16 1.17 0.16 0.22 1.47 0.14 Th e sample contained about: 2% Pb , 2 oz/ton Ag , 3% Zn, and 1 0% Fe . The preliminary test conditions and results are outlined in Table 13 above. In evaluating the above results, the fact that this was a "blind test" is entitled to substantial weight. The above results may be used to estimate those of an industrial scale application in regular operation, by extrapolation. Further laboratory testing could be done to further reduce the amount of pyrite collected with the zinc rougher concentrate. The above results indicate excessive activation by CuSO., which may be controlled by exercise of ordinary skill in the art. ORE F: Sample from run of mine mixed sulfides containing approximately 0.18% Pb, 8.4%Zn and 10-12% FeS 2 by weight. The testing procedure involved wet grinding to 85% passing 65 mesh. The reagents used, testing procedure and results are summarized in Tables 14-17, below, and show substantial recoveries and selectivity. OMPI & . TABLE 14 % Distribution Weight PRODUCT % % Pb % Zn Pb Zn Pb Ro Con ( 1 ) 4.65 .63 2.28 69.29 1.26 Zn Ro Con (2) 10.40 .10 61.37 5.89 75.93 Zn Sc 1 Con (3) 2.90 .13 48.28 2.14 16.68 Zn PRIM Con (1-3) 17.19 .12 46.83 11.33 95.76 Zn Sc 2 Con 3.89 .15 6.82 3.30 3.15 c FeS 2 Ro Con 9.93 .07 .73 3.94 .86 00 NON-FLOAT 68.23 .04 .26 15.45 2.11 HI HEADS 100.00 .176 8.40 100.00 100.00 Α c Time STAGE (Min.) £H REAGENTS (G/T) m Grind 8 7.65 50g/ton a 2 S ω x Cond. I 5 50g/ton NaCN PI Pb Cond./Flot. 5/5 20g/ton A-242, 15g/ton frother Cond. II 4 7.5 150g/ton CuSO. Zn Rougher Cond.Flot. 5/2 30g/ton Z-14 Zn SC- j F iot. 3 TABLE 15 % Distribution Weight PRODUCT % % Pb % Zn Pb Zn Pb Ro Cone. 4.71 2.55 2.43 71.43 1.26 Zn Ro Cone. (1 ) 12.43 .10 59.15 7.39 87.54 CA Zn Sc. Cone. (2) 1.93 .21 25.18 2.41 5.79 G CD Zn Prim. Conc(1-2) 14.36 .11 54.60 9.80 93.33 21 FeS 2 Ro Cone. 12.19 .09 1.46 6.52 2.12 c NON-FLOAT 68.73 .03 .39 12.25 3.19 HEADS 100.00 .168 8.40 100.00 100.00 rπ Cond. I 5 — 75g/ton NaCN Pb Cond./Flot. 5/5 — 15g/ton A-242, 15g/ton frother Cond. II 7.5 200g/ton CuS0 4 Zn Rougher Cond. Flot. 5/2 30g/ton Z-14 Zn Sc. Flot. 3 TABLE 16 % Distribution Weight PRODUCT % % Pb % : Zn Pb Zn Pb Ro Cone. 4.28 2.82 2.03 69.52 1.03 Zn Ro Cone. 14.58 .10 51 .67 8.39 89.59 Zn Sc Cone. 3.12 .13 12 .52 2.33 4.64 Zn Prim. Cone. 17.70 .11 44 .77 10.72 94.23 FeS 2 Ro Cone. 7.84 .08 1, .50 3.61 1.40 NON-FLOAT 70.17 .04 .40 16.15 3.34 HEADS 100.00 .174 8, .41 100.00 100.00 STAGE Time Eϋ REAGENTS (GR/MT) Grind 8' — Cond. I 5' — 100g/ton a 2 S Cond. II 5' — 100g/ton NaCN Pb Cond./Flot. 5*/5' 20g/ton A-242, 15g/ton frother Cond. Ill 5' 7.5 200g/ton CuS0 4 Zn Rougher 5 f /2 f — (15g/ton frother). Cond./Plot. 50g/ton Z-14 Zn Sc. Flot. 3' — — FeS 2 Ro 3'/5* — 1 5g/5on frother. Cond/Flot. 50g/ton Z-6 TABLE 17 % Distribution Weight PRODUCT % % Pb % Zn Pb Zn Pb Ro Cone. 5.61 2.48 3.05 72.03 2.14 Zn Ro Cone. 12.87 .07 55.38 4.67 89.40 Zn Sc Cone. 4.23 .20 2.88 4.38 1.53 Zn Prim. Cone. 17.10 .10 42.40 9.05 90.93 FeS 2 RO. Cone. 9.36 .10 4..09 4.85 4.80 NON-FLOAT 67.94 .04 .25 14.08 2.13 HEADS 100.00 .193 7.97 100.00 100.00 Time STAGE (Min. ) £H REAGENTS (GR/MT) Grind 8 100g/ton Na 2 S Pb Cond./Flot 5/5 7.5 20g/ton A-242, 15g/ton frother Condit. 3 10.9 1330g/ton LIME Zn Rougher Cond./Flot. 5/2 10.7 (15 g/ton frother) 465g/ton Cu S0 4 , 50 g/ton Z-14 Zn Sc. Flot. 3 7.5 g/ton frother FeS 2 Ro Cond/Flot. 3/5 — 15g/ton frother, 50g/ton Z-6 Table 17 presents test results obtained with use of lime and is set forth above for comparison purposes. ORE G: Zinc Dumps processed at Don Diego, Potosi, Bolivia containing 35% sphalerite and 20% pyrite. Treated in accordance with Fig. 1. The natural ore pH was 5.5. ' TABLE 18 Weight % Dist. (Tons) % Zn Zn Day 1 Feed 143.65 8.02 100.00 Conct. 40.65 56.57 88.85 Tail 103.00 2.80 11.15 Day 2 Feed 114.88 18.40 100.00 Conct. 34.54 56.09 91.67 Tail 80.34 2.19 8.33 Day 3 Feed 95.71 18.79 100.00 Conct. 31.74 53.73 94.81 Tail 63.97 1.46 5.19 Reageiit Consumption: Na 2 S 75 g/t; NaCN 149 g/t; CuS0 4 1088 g/t; Z-200 75 g/t; Z-6 103 g/t; Frothers 34 g/t OMPI The particular applications of the present in¬ vention to concentration of Cu-Mo are further illustrated by the following additional examples: ORE H: Sample consisting of pyrite, molybdenite, chalcopyrite and chalcocite finely dispersed in quartz monzonite porphyry. Run of mine ore was ground to 80%-100 mesh* (Tyler) during all tests following operating plant proced¬ ures. The first two tests (results and conditions set forth in Tables 18-19) involved induced flotation in accordance with Fig. 2, one without lime, one with lime. The last two tests (results and conditions set forth in Tables 20-22) involved collectorless flotation according to Fig. 3 using a combination of Na 2 S and NaCN. Collec¬ torless flotation using the present invention gave a Mo rougher concentrate of a better grade. Finally, Table 23 summarizes collectorless flotation without use of NaCN (for comparison purposes). Table 23 shows better Mo-Cu separation but poorer Cu-pyrite separation. TABLE 18 % Distribution Weight Analysis % PRODUCT % Mo Cu Mo Cu Moly Ro. Cone. 2.37 5.00 3.89 80.22 60.70 Copper Ro. Cone. 2.08 .42 .79 5.91 10.82 Pyrite Ro. Cone. 1.63 .45 .81 4.97 8.69 Non-Float 93.92 .014 .032 8.90 19.79 Heads 100.00 .152 .148 100.00 100.00 Time STAGE (Min.) pH REAGENTS (g/ton) Grind 5.5 . — 50 Na 2 S, 100 Moly-Copper Collector, Cond. I 5 7.3 Mo. Ro. Flot. 5 7.9 100 Na 2 Si0 3 , 75 NaCN, 15 frother (MIBC) Cu.Ro. (Cond./ Flot.) 5/5 7.5 frother (MIBC), 5(1331)** Pyrite Ro. 3/5 7.5 frother (MIBC) 50(Z-6) (Cond./Flot) OMPI Footnotes to p. 29: * At the given grind size, liberation of only 80% of each Cu and Mo was obtained. ** MINEREC 1331 (copper collector). TABLE 19 % Distribution Weight Analysis % PRODUCT % Mo Cu Mo Cu Mo-Cu Ro. Cone. 2.9 3.73 2.85 78.64 56.05 Mo-Cu Scav.Conc. 1.09 1.12 .77 8.84 5.67 Non-Float 96.00 .018 .059 12.52 38.28 Heads 100.00 .138 .148 100.00 100.00 Time STAGE (Min.) £H REAGENTS (g/ton) Grind 5.5 9.5 1000 Lime ), 100 MCO Collector* Cond. I 10.7 500 (Lime-)-5(1331) 7.5 (MIBC) Mo-Cu Ro Flot. 5 Mo-Cu Scav. Flot. 5 * Mo-Cu Collector (Phi .Hips 66 Co.) TABLE 20 % Distribution Weight Analyεsis % PRODUCT % Mo Cu Mo Cu Mo Ro. Cone. 1.48 3.52 2.63 54.36 30.47 Mo. Scav. Cone. 1.00 .98 1.92 10.25 15.07 Cu Ro. Cone. 1.50 .46 1.79 7.20 20.54 Cu Scav. Cone. 1.07 .38 .62 4.25 5.2 FeS 2 Ro Cone. 2.15 .16 .54 3.59 9.1 Non-Float 92.80 .021 .027 20.35 19.63 Heads 100.00 .096 .128 100.00 100.00 y tΕA. I STAGE Time REAGENTS (g/ton) Grind 5.5 7.9 50(Na 2 S) Cond. I 3 50(Na 2 S) Cond . I I 3 25 (NaCN), 15(frother) Mo. Ro. Flot. 5 Mo. Scav. Flot. 5/5 7.5 (frother), 10 (fuel oil) Cu Ro. Cond. Flot. 3/5 15 (frother), 5/Z-14) Cu Scav. Cond. Flot. 3/5 7.5 (frother), 5(Z-14) Pyrite Ro Flot 3/5 15 (frother, 25 (Z-6) • TABLE 21 % Distribution Weight Analysis ; % PRODUCT % Mo Cu Mo Cu Mo. Ro. Cone. 2.24 3.47 2.30 69.05 42.53 Mo. Scav. Cone. .89 .93 .86 7.34 6.30 Cu Ro. Cone. 2.59 .21 1.28 4.82 27.32 Pyrite Ro. Cone. .89 .28 .31 2.21 2.27 Non-Float 93.40 .02 .028 16.59 21.58 Heads 100.00 .113 .121 100.00 100.00 STAGE Time EH REAGENTS (g/t<on) Grind 5.5 8.1 150ι (Na 2 S) Cond. I 3 50 (Na 2 S) Cond. II 3 25 (NaCN) Mo. Ro. Flot. 5 Mo. Scav. Flot. 5/5 7 .5 (frother) ,10 (fuel oil) Copper Ro. Cond.Flot. 3/5 15 (frother) , 10 (Z-14) Pyrite Ro. Cond.Flot. 3/5 15 (frother) , 25 (Z-6) TABLE 22 % Distribution Weight Analysis % PRODUCT % Mo Cu Mo Cu Mo. Ro. Cone. 1.65 S.66 2.12 59.41 27.96 Mo. Scav. Cone. .89 .99 2.20 8.61 15.55 Copper Ro. Cone. 1.36 .48 1.74 6.40 18.86 Copper Sc. Cone. .54 .46 .83 2.44 3.59 Pyrite Ro. Cone. 2.36 .13 .70 3.01 13.20 Non-Float 93.20 .022 .028 20.13 20.83 Heads 100.00 .102 .125 100.00 100.00 rrime STAGE (Min.) pH REAGENTS (g/ton) Grind 5.5 7.9 75 (Na 2 S) Cond. I 3 25 (Na 2 S) Cond. II 3 25 (NaCN), 15 (frother) Mo. Ro. Flot. 5 Mo. Scav. Cond.Flot. 5/5 7. .5 (frother), 10 (fuel oil) Copper Ro.Cond.Flot. 3/5 15i (frother), 5 (Z-14) Copper Sc.Cond.Flot. 3/5 7. 5 (frother), 5i (Z-14) Pyrite Ro.Cond.Flot. 3/5 15i (frother), 2Ξi (Z-6) TABLE 23 % Distribution Weight Analysis % PRODUCT % Mo Cu Mo Cu Mo Rougher Cone. .98 9.25 .64 72.65 6.00 Mo. Scav. Cone. .55 1.46 .65 6.47 3.47 Copper Ro. Cone. .69 .32 1.45 1.76 9.56 Copper Sc. Cone. 1.10 .42 .82 3.71 8.67 Pyrite Ro. Cone. 2.04 .11 2.22 1.79 43.34 Non-Float 94.63 .018 .032 : 13.61 28.97 Heads 100.00 .125 .105 100.00 100.00 Time STAGE (Min.) pH REAGENTS (q/ton) 10x Grind 5.5 60 (Na 2 S) Cond. I 3 20 (Na 2 S), 7.5 Mo. Ro. Flot. 7.5 7.4 (frother) Mo. Scav. Cond. Flot. 5/7.5 2.5 (frother), 7 (fuel oil) Copper Ro.Cond. Flot. 3/10 5 (Z-14) Copper Sc.Cond. Flot. 3/5 2.5 (frother), 2 (Z-14) Pyrite Ro.Cond. Flot. 3/5 30 (Z-6) Theoretical Calculation In a typical concentration of Cu-Mo containing ore in accordance with the prior art treating 20,000 tpd of 0.7% Cu and 0.015% Mo, primary flotation will produce OMPI 476 tpd of a bulk Cu-Mo concentrate assaying 25% Cu and 0.536% Mo, representing a Mo recovery of 85%. A primary flotation process in accordance with Fig. 3, with the same recovery would only have to produce 85 tpd of a molybdenite float assaying 3% Mo and 3% Cu. In addition, this 85 tpd would be essentially collector-free, thus eliminating the need for collector removal or transformation. OMPI y/Λ,- WIPO

Claims: What is claimed is: 1. A process for the separation of ore compon¬ ents by flotation comprising: grinding ore to form ore pulp-, mixing said pulp with sulfide ions and cyanide ions, adjusting the concentration of said sulfide ions to a level at least sufficient to cause depression of base metal mixed sulfides but insufficient to cause substantial activation of pyrites; and adjusting the concentration of said cyanide ions to a level at least sufficient to obtain auxiliary depression of the mineral components of said ore which are required to be depressed in said flotation, but insufficient to cause overde- pression of said mineral components; said sulfide ions and cyanide ions having been introduced to said pulp at a prede¬ termined time and in a predetermined sequence, prior to flotation. 2. The process of claim 1, wherein said sulfide ions are introduced during grinding of said ore, said grinding being wet grinding. 3. the process of claim 2, wherein said cyanide ions are introduced into said pulp subsequent to the introduc¬ tion of said sulfide ions. 4. The process of claim 3, wherein said cyanide ions are introduced into said pulp in a conditioning operation following said wet grinding. 5. The process of any one of claim 1-4, said process taking place at a substantially unmodified pH. 6. The process of any one of claims 1-4, said process taking place at an unmodified pH. 7. The process of any one of claims 1 - 4 wherein said ore is a complex base metal mixed sulfide ore containing at least two of the metals Pb, Cu, Ag, Zn, Fe, the process further comprising subsequent conditioning of said pulp with collectors and frothers followed by direct selective flotation of the valuable mineral constituents of said ore in the order of: Pb-[Ag] : Cu : Zn : Fe. 8. The process of claim 3 or 4, wherein said ore is a Cu-Mo ore, the process further comprising subse¬ quent conditioning of said pulp with a copper collector and frother followed by flotation of a Cu-Mo concentrate. 9. The process of claim 3 or 4, wherein said ore is a Cu-Mo ore, the process further comprising subse¬ quent direct collectorless flotation of a Cu-Mo concentrate. 10. The process of any one of claims 1 - 4 wherein said ore is coal, the process further comprising introducing frothers and collectors in said pulp, and floating said coal while said sulfides remain in the gangue. 11. The process of claim 7, wherein said sulfide ion is provided by a member selected from the group consisting of Na 2 S, K 2 S and NaHS. 12. The process of claim 11, wherein said cyanide ion is provided by a member selected from the group consisting of NaCN, KCN and Ca(CN) 2 . 13. The process of claim 12 wherein said sulfide ion is provided by Na 2 S. 14. The process of claim 13 wherein said cyanide ion is provided by NaCN. 15. The process of claim 13 or 14 wherein said Na 2 s consumption ranges between about 20 and 200 g/ton. 16. The process of claim 15 wherein said NaCN con¬ sumption ranges between about 25 and 200 g/ton. 17. The process of claim 8 wherein said sulfide ion is provided by a member selected from the group consisting of Na 2 S, K 2 S and NaHS. 18. The process of claim 9 wherein said sulfide ion is provided by a member selected from the group consisting of Na 2 S, K 2 S and NaHS. 19. The process of claim 17 or 18 wherein said cyanide ion is provided by a member selected from the group consisting of NaCN, KCN and Ca(CN) 2 20. The process of claim 17 or 18 wherein said sulfide ion is provided by Na 2 S. 21. The process of claim 20 wherein said cyanide ion is provided by NaCN. 22. The process of claim 20 wherein said Na-S consumption ranges between about 20 and 50 g/ton. 23. The process of claim 21 wherein said NaCN con¬ sumption ranges between about 25 and 100 g/ton. 24. The process of claim 8, said process taking place at a substantially unmodified pH. 25. The process of claim 9, said process taking place at a substantially unmodified pH. 26. The process of claim 7, said process taking place at a substantially unmodified pH. 27. The process of claim 10, said process taking place at a substantially unmodified pH. 28. The process of any one of claims 24 - 27, said process taking place at an unmodified pH. _ OMPI _ 29. A process for the separation of ore components by flotation comprising: grinding ore to form ore pulp, mixing said pulp with sulfide ions and cyanide ions at an effective concentration, said sulfide ions an yanide ions having been introduced to said pulp at predetermined times and in a prede¬ termined sequence, prior to flotation, whereby at least one of the following is effected: reagent quantities are reduced, flotation selectivity is improved, recovery is increased, concentrate grades are improved and conditioning and residence times are reduced. OMPI °

Inventors: VARGAS ALFREDO PERCY, VARGAS, ALFREDO, PERCY
Applicants: PHLOTEC SERVICES INC, PHLOTEC SERVICES INC.
Application year: 1983
Main classification: B03B1
IPC codes: B03B1, B03D1