US3635337A - Method for treating floated solids - Google Patents

Abstract

To prevent the flotation of a mineral containing an anionic collector reagent, the coated mineral is formed into a thick alkaline pulp containing another mineral having a greater affinity for the collector and an alkaline dispersant at a high concentration. The pulp is aged and then diluted and aerated. The mineral having the greater affinity for the anionic reagent floats; the other mineral is depressed.

Description

United States Patent Mercade et al.

[451 *Jan. 18, 1972 METHOD FOR TREATING FLOATED SOLIDS lnventors: Venacio Mercade, Metuchen; Samuel R.

Weir, Long Branch, both of NJ.

Engelhard Minerals & Chemicals Corporation, Township of Woodbridge, NJ.

Assignee:

Notice: The portion of the term of this patent subsequent to July 8, 1986, has been disclaimed.

Filed: Aug. 23, 1968 Appl. No.: 754,952

US. Cl ..209/3, 209/166 Int. Cl. ...B03b l/00, B03d 'l/06 Field of Search ..209/5, 166, 167, ll, 3

[56] References Cited UNITED STATES PATENTS 1,447,973 3/1923 Feldenheimer ..209/5 3,331,505 7/1967 Mercade ....209/l1 3,454,161 7/1969 Mercade ..209/167 Primary Examiner-Frank W. Lutter Assistant Examiner-Robert Halper Attorney-Melvin C. Flint ABSTRACT 10 Claims, No Drawings METHOD FOR TREATING FLOATED SOLIDS This invention deals with an improved method for depressing particulate solid material containing an anionic collector reagent. The invention is especially directed to the flotation art and to the separation of minerals in slimed bulk flotation concentrates obtained by negative-ion flotation concentration. More particularly, the process of the invention represents an improvement over practices set forth in U.S.

Pat. No. 3,331,505 to Venancio Mercade.

BACKGRbUND OFTHE INVENTION The differential flotation of a mineral from other mineral matter having similar flotation characteristicsis especially difficult when the minerals are slimed. Fatty acid collectors, for example, will float various oxide minerals from silica and silicates, especially when well-dispersed pulps are used. However, the collector generally is not effective in selectively separating slimed oxidized mineral from each other. Similarly, bulk'flotation of slimed sulfide minerals may be made with fatty acids but selectivity is normally not realized.

it has been suggested in U.S. Pat. No. 3,331,505 to Venancio Mercade to effect the differential flotation of a slimed calcareous mineral, especially calcite, from a slimed noncalcare-' ous mineral in a fatty acid reagentized bulk float product by forming the bulk float into a dilute pulp (.10 to 30 percent solids) containing an alkaline dispersant such as sodium silicate at a concentration of l to g./l. The dilute pulp is agitated until the froth is clear and nonslimy (about 10 minutes) while maintaining the pulp at a high temperature, e.g., 180 F. After the froth or foam on the surface of the pulp becomes crystal clear, air is introduced under the surface of the pulp. As a result, the calcareous mineral floats and is separated from the noncalcareous mineral which remains dispersed in the pulp. Apparently reagent is transferred from the noncalcareous mineral to the calcite by the treatment with the hot dispersant solution. This may be attributed to the fact that the collection of calcite is facilitated by the use of high temperature. At any rate, under conditions of treatment employed in accordance with the teachings of the patent, heat is essential during the dispersant treatment.

While unique separations have been effected by the process described in the Mercade patent, the dispersant treatment of the bulk float to render the minerals separable by flotation without addition of another collector is undesirably costly for some beneficiation treatments. The pulps are dilute, requiring a large volume of space and equipment. The power and heating requirements add further to the processing costs.

THE lNVENTlON An object of the invention is to provide an improved process for treating a fatty acid reagentized bulk float product with a dispersant (deflocculating agent) so that the solids in the bulk float product may be separated by flotation without addition of a collector reagent.

Another object is to provide an improved process for treating a solid containing a fatty acid collector reagent so that the solid is no longer floatable.

Another object is to provide a novel method for separating slimed minerals having similar flotation characteristics.

Further objects and features will be apparent from the description which follows.

Briefly stated, in accordance with this invention, a semisolid or solid aqueous mixture of a particulate solid coated with an anionic reagent and a different particulate solid having a greater affinity for said reagent is thinned by adding an a] kaline dispersing (deflocculating) agent without diluting the mixture to a solids content below 50 percent by weight. The

in accordance with an embodiment of the invention, the solid having the greater avidity for the reagent may already be reagentized with the anionic collector when it is formed into the pulp, e.g., the solids in the semisolid to solid mass may consist of a dewatered bulk float product obtained with an anionic collector. Metallurgical results not heretofore considered possible have been realized when processing lowgrade slimed ore pulps by a primary bulk flotation with an anionic collector, followed by dewaterlng of the bulk float, addition of dispersant, aging, dilution-and aeration, all in accordance with this invention. By carrying out the bulk float before the secondary differential step, greater recoveries and concentration of valued minerals may be realized than when attempts are made to recover the slimed mineral value initially by itself with a selective reagent.

Pursuant to another embodiment, the bulk float is a slimed ultraflotation" froth concentrate obtained by the procedure described for example in U.S. Pat. No. 2,990,958 to Ernest W. Greene et al. The process features the use of a reagentized finely divided solid additive such as calcite in a slimed fatty acid reagentized ore pulp. The added oiled solid aids in the collection of the slimed selectively reagentized constituent of the ore and the two report in a bulk float concentrate. By applying the process to such a bulk float concentrate, the carrier is separated from the ore constituent, usually as the float product in the differential flotation step. The carrier may be reused in subsequent ultraflotation after being reoiled. The ore constituent thus separated may be'further processed.

The mineral having the greater affinity for the anionic reagent may be one that is added in unreagentized condition with a previously oiled solid for the purpose of depressing the solid.

PRlOR ART v To the best of our knowledge, aging of high-solids reagentized pulps in concentrated dispersant solutions has never been employed for any purpose, much less to permit subsequent differential flotation without addition of selective reagents. We are aware that it has been suggested in U.S. Pat. No. 2,811,254 to McGarry to age bulkphosphate-silica floats at high solids. However, the aging treatment is not carried out in the presence of an added reagent such as our dispersant. To the contrary, mineral acid, the reagent that removes the anionic collector in the McGarry process, is added after the aging step and it is not present during the aging treatment. Moreover, the anionic reagent is removed on a nonselective basis in the McGarry process. Addition of a different type of collector (cationic) is then required to effect the differential flotation in the secondary flotation treatment. ln contrast, our aging process, carried out in the presence of added dispersant, permits differential flotation without addition of a cationic collector since our dispersant treatment results in selective depression.

The process of the invention represents a substantial improvement and departure from practices set forth in said Mercade patent. A fraction of the equipment and space is required for the dispersant treatment and the, agitating and heating equipment may be dispensed with. For example, when the dispersant treatment of the present invention is carried out at a preferred pulp solids level, e.g., 75 percent, the process requires less than half the volume of equipment that would be used with the process of the patent at a maximum solids of 30 percent. With the high-solids pulps we employ during the aging treatment, the dispersant is present at a high concentration. When employed in quantity suflicient to thin the aqueous mixtures-as a result of deflocculation, the minimum concentration of dispersant is at least 50 percent greater than the maximum concentration of dispersant employed in carrying out the Mercade process. in most applications of our process, the dispersant concentration is at least 500 percent greater than the maximum dispersant concentration in the Mercade process.

It was surprising and unexpected that selective flotation could be carried out after a pulp had been treated at the highdispersant concentrations employed in our process. The prior art contains numerous references to the fact that minerals are depressed by alkaline dispersants when the dispersants are employed at concentrations that are significantly lower than those we employ. It would have been expected that all minerals including those previously collected would have been depressed by the treatment and that differential flotation would therefore not be possible. The beneficial results of the high-solids treatment at high-dispersant concentrations were not predictable from those realized in carrying out the process of the Mercade patent. The effect of high temperature on the collector action of a fatty acid which might explain the operativeness of the Mercade process would not suffice to explain why selectivity could be realized in carrying out our process which does not require the use of elevated temperatures. It is significant that under the conditions of solids and reagent con centrations described in the Mercade patent differential separation was not realized at temperatures that may be employed in carrying out the process of this invention.

DETAILED DESCRIPTION The process is broadly applicable to the treatment of mineral and/or nonmineral solids and is of special importance and value in the treatment of slimed, e.g., minus 200 mesh (Tyler) solid matter.

The anionic collector may be any carboxy-containing reagent such as a fatty acid or soap, especially a mixture of fatty and resin acids (or soaps thereof). As is known in the art, anionic collectors usually also include hydrocarbon oils such as fuel oil and frequently also contain emulsifying agents such as oil-soluble petroleum sulfonates. Examples of anionic collectors include oleic acid, tall oil acids (refined or crude), sulfo-oleic acid, soaps thereof, mixtures of the aforementioned with each other or with hydrocarbon liquids such as fuel oil and/or petroleum sulfonate salts. Especially good differential separations have been made from bulk floats containing a collector mixture of tall oil acids and an oil-soluble petroleum sulfonate.

The process of the invention is of especial importance in connection with differential flotation of minerals in slimed, flocculable-deflocculable bulk float concentrates containing anionic collectors. In these processes, at least two mineral species are reagentized with an anionic collector or collector mixture. By applying our process to such concentrates, the mineral species are separated and recovered without the need to remove reagent from the bulk float as a whole and then add a different type of collector to effect the differential flotation. Separations not heretofore considered possible have been accomplished when the process of the invention was applied to low-grade complex oxidized and sulfide ores. Especially good results have been realized separating slimed oxides of transition metals (e.g., titanium dioxide such as anatase, tin oxide such as cassiterite, and manganese dioxide such as pyrolusite) from calcite in finely disseminated ores. After the bulk float was taken and the float product was dewatered, dispersed, aged and aerated, the calcite floated and the transition metal oxide was recovered in the flotation tailings. For example, manganese dioxide (pyrolusite) was recovered from a finely disseminated ore containing calcite and siliceous gangue by producing a bulk float with a fatty acid and then separating calcite by flotation from the pyrolusite by treating the bulk float in accordance with this invention. The separation and recovery of pyrolusite by this method is claimed in a copending application, Ser. No. 754,951, filed of even date herewith. Zinc sulfide and tin sulfide minerals have also been separated from each other and gangue by similar treatment. The process is claimed in a copending application, Ser. No. 718,324, filed Apr. 3, l968 now U.S. Pat. No. 3,454,161.

The present invention is also of value in the treatment of Ultraflotation froth products resulting from the flotation of slimed metal oxide minerals, especially slimed anhydrous tetravalent metallic elements (such as anatase and cassiterite), with oiling (collector-coating) reagents including fatty acids and carrier particles which also have an affinity for fatty acids and are rendered floatable thereby. Minerals with surface characteristics similar to the aforementioned tetravalent metal oxides are rutile and zircon. Other metal oxides include iron oxides and manganese oxides.

The process of the invention is applicable to the treatment of Ultraflotation froth products containing various types of auxiliary solid or carrier flotation particles. Generally speaking, the auxiliary solid flotation reagents are characterized by being very finely divided (minus 325 mesh) homogeneous particles that are different in composition from the slimed pulp to be conditioned therewith. The carrier particles are oiled by fatty acids and the oiled particles are floatable in the conditioned pulp. High-purity minerals, exemplified by minus 325 mesh alkaline earth carbonate minerals, constitute one class of carrier reagent and are preferred in Ultraflotation processes for reasons of economy. Alkaline earth carbonate particles include calcium carbonate, such as ground marble or calcite, magnesium carbonate, barium carbonate or mixed calciummagnesium carbonates. Nonwaxy organic solids, especially nonwaxy polymeric solids, can also be used in Ultraflotation. Examples of plastic carriers are polyethylene, polypropylene, polystyrene, polyvinyl chloride, polyamides and the like. Natural floaters, such as talc or the like, can also be used. Sulfur and fluorite are examples of other carrier solids, Especially good results are obtained with micron-size or slimed carrier particles, e.g., particles finer than 10 microns.

The Ultraflotation bulk concentrates are obtained by the selective flotation of mineral particles from ore pulps fine enough to pass 200 mesh (Tyler) screens. in some cases the pulps are fine enough to pass a 325 mesh screen. The froths can also be obtained from the Ultraflotation concentration of slimed pulps containing some coarse (e.g., plus 65 mesh) particles.

ln carrying out the process of the invention, froths obtained by bulk flotation with negative-ion collector are thickened or dewatered, preferably by filtration, producing cakes of at least 50 percent solids. It is preferable to filter the cakes until the solids are within the range of 70 to percent since the use of high-solids cakes permits the formation of high-solids pulps which generally produce better results. There is no upper limit to the solids content of the cakes except for the practical limitations involved in draining all of the water from the cakes. Since the water must be subsequently added to the filtered froth to dissolve the dispersing agent, it would be impractical to incur the expense of removing all water. While the froths can be dried by the use of heat, the drying must be limited to the use of conditions which do not result in the destruction of the oils in the froths or in the cementation of the components. Decantation followed by filtration, or filtration alone, are preferred methods for dewatering since they result in the removal of pulp water which contains reagents that may interfere with subsequent operations. Also, when pulp water is removed by decantation or filtration (alone or in combination with each other), entrained water-avid solids in the bulk float products may be removed simultaneously.

Solids contents are determined by weighing the material before and after oven drying at l75 F. for 1 hour and calculating as follows:

Percent solids Weight (original) weight (after drying) 7 W i a The defiocculating agent can be added to the dewatered pulp or filter cake as dry reagent or as a concentrated solution. Incorporation of the defiocculating agent to the solid or semisolid filter cake results in a distinct thinning or fluidization of the cake. Normally the fluidized mass is very viscous and is frequently distinctly thixotropic. Since oiling reagents are also present, the thinned cake usually has a creamy appearance, similar to that of a milk shake. As mentioned above, the water content of the deflocculated pulp must be restricted. Therefore, any water added to the filtered froth, separately or with the defiocculating agent, must be limited. The addition to the filter cake of the defiocculating agent as a concentrated solution results in a very small decrease in solids level.

To effect the deflocculation and resulting thinning of the flocculable-defiocculable solid or semisolid reagentized mass, the dispersant is added to the fiocculated semisolid mass in amount within the range of about to 50 lb./ton of solids (anhydrous dispersant basis). When this quantity of dispersant is added without reducing the pulp solids level below about 50 percent, as required in our process, the dispersant concentration will be at least grams per liter of water in the pulp; With most pulps, a dispersant concentration of at least 25 grams per liter is required. Especially preferred is a dispersant concentration of at least 50 grams per liter. By way of example, when 0 sodium silicate solution (38 percent solids) is added without dilution to a 73 percent solids flocculabledeflocculable aqueous pulp, the sodium silicate concentration (anhydrous basis) will be about 55 grams per liter of water in the pulp.

Stage addition of the defiocculating agent may be practiced, if desired.

After the defiocculating agent has been added to the dewatered froth, the resulting thick or viscous creamy pulp must be aged to permit the defiocculating agent to act upon the constituents of the pulp in the desired manner. During the aging, the pulp should contain at least 50 percent solids, preferably 60 percent solids or more. Especially preferred is aging at a solids content within the range of 70 to 80 percent by weight. The rate at which the defiocculating agent acts varies directly with temperature. When the aging is carried out at ambient temperature, times of at least 12 hours, preferably at least 18 hours, are required. Aging can be carried out at temperatures up to about 120 F. if desired, with corresponding reduction in time. There is no apparent advantage in our process to the use of very hot pulps. Prolonged aging is not detrimental.

The thick pulp may be maintained quiescent during aging or it may be agitated if sticking of the pulp to the sides of the vessel is a problem. Opened or closed vessels may be used during the aging treatment.

After the pulp has aged a suitable amount of time at the high-solids level, the pulp is diluted before the aeration and separationis carried out. Dilution of a slimed pulp to a solids level of about 3 to 30 percent, preferably about 5 to 10 percent, is recommended. Higher solids may be used when coarser pulps are employed. Dilution of the pulp results in a decrease in the concentration of dispersingagent; additional dispersants or collectors are not added.

The diluted pulp is subjected to aeration in a suitable flotation cell. Aeration results in the flotation of oiled particles. Usually the carrier particles float when a bulk Ultraflotation concentrate is treated. When reagent-avid solid such as calcite or fluorite has been added to depress a solid coated with an anionic collector, the former solid is the one that floats. In the case of a bulk float consisting only of ore minerals, the float product is a concentrate of the mineral that has the greater affinity for the collector. For example, calcite is concentrated in the float product when the bulk float has been obtained by anionic flotation of an ore pulp containing calcium carbonate and an oxide of a transition metal such as anatase, pyrolusite, cassiterite. The mineral (or minerals) having the lesser affinity for the reagent or reagent mixture remains dispersed and depressed in the flotation cell.

The float products and tailings may be further processed by appropriate physical and/or chemical means. Minus 325 mesh calcite is a preferred solid when it is desired to depress a fatty acid reagentized oxide or sulfide mineral by the process of the invention. Other solids may be employed. The dispersant and solid additive are added to a dewatered float product containing the mineral to be depressed, using sufficient water and dispersant to provide a total solids content of at least 50 percent and a suitable dispersant concentration, as described above.

The invention and some of its features and benefits will be better understood from the following illustrative examples.

In the examples which follow all reagents are reported on an anhydrous weight basis unless otherwise indicated. Water that had been deionized by treatment with cation-exchange resins and anion-exchange resins was used throughout the mineral processing steps.

Examples l to VII illustrate the application of the process of the invention to the separation and recovery of a carrier" mineral in a flotation process. The carrier used in these tests was minus 325 mesh calcite (Drikalite) obtained by grinding marble and classifying it to a mean particle size below 5 microns. The Drikalite" is substantially pure calcium carbonate and has a GE. brightness value of 91 percent as supplied. This carbonate mineral was used as the carrier in the U1- traflotation concentration of colored (yellow-brown) anatase from kaolin clay with a fatty acid collector reagent in an alkaline pulp. As a result of the presence of the anatase in the froth product of the Ultraflotation beneficiation of the clay, the brightness of the Drikalite was reduced. Since the brightness of the treated froth generally varied inversely with anatase content, the brightness value of the treated froth was used in some instances to estimate purity of the calcite float product.

EXAMPLE I The froth product that was used in the test was obtained from a commercial kaolin flotation operation, substantially as described in an article by Frank A. Seeton, Ultraflotation," Bulletin No. M4-B117, Denver Equipment Company. The clay charge was a Georgia kaolin from a mine near McIntyre, Ga. The oiling agents employed included tall oil and a solution of neutral oil-soluble petroleum sulfonate in mineral oil (Calcium Petronate). Fuel oil has been employed during the flotation to control the froth consistency. As supplied, the froth contained about 40 percent solids and assayed 6.78 percent TiO The froth had a mustardlike appearance and was composed of a mixture of the Drikalite carrier, colored anatase, adherent clay and oiling reagents.

The fresh froth from the flotation cell was washed free of clay by the following steps. 0" sodium silicate solution was added to the froth in amount of 3 lb./ton solids and the mixture was diluted with deionized water to 1 percent solids. 0 sodium silicate solution contains 38 percent solids and has a Na O: SiO mole ratio of 123.22.) The pulp was main tained quiescent to permit sedimentation. As a result, the clay formed a dilute clay-water suspension. Finer portion of the oiled froth solids floated on the surface of the suspension. The coarser flocs in the froth settled to the bottom of the claywater suspension. The clay-water suspension was removed from the nonclay solids by siphoning the suspension. The solids were again diluted to 1 percent and washed by sedimentation and siphoning. The procedure was repeated until the froth was washed three times.

The froth solids remaining after the clay was removed were filtered, producing a solid filter cake containing 65 percent solids.

Q sodium silicate solution was added to the filter cake in amount of 1.3 grams, corresponding to 26 lb./ton of cake solids, followed by 0.515 grams dry sodium carbonate, corresponding to the use of 13 1b./ ton of this reagent. The dispersant concentration was calculated to be 55.7 g./l. (anhydrous basis). The filter cake was deflocculated by the addition of the sodium silicate and sodium carbonate, producing a viscous, thixotropic mass having the appearance of a thick milk shake. The deflocculated filter cake was agitated by a magnetic stirrer at low speed for 18 hours at ambient temperature. Since the dispersants were added in concentrated form, the solids content of the dispersed filter cake was only slightly less than 65 percent in spite of the fact that the dispersant had been employed as a solution.

After being stirred for 18 hours, the thick mobile mass was diluted to percent by adding deionized water and the diluted pulp was placed in a l,000 g. capacity AirFlow flotation cell. The charge in the cell was aerated and a float removed. The float was cleaned twice by flotation without addition or reagents.

The final float product was the calcite concentrate. This concentrate was dried, weighed and analyzed for TiO The TAPPl procedure was used to measure G.E. brightness.

The three machine discharges were combined to form the anatase concentrate. This concentrate was flocced by adding alum and the flocced product was dewatered by decantation, dried and weighed. To determine the TiO content, residual carbonate was destroyed by adding a 10 percent solution of hydrochloric acid to the weighed material. The residue was filtered, washed and analyzed for TiO The results are summarized in table I.

TABLE I.-FLOTATION OF CALCITE FROM ANATASEIN CLAY FLOTATION FROTH CONCENTRATE l Calculated using fact that feed contained 90.1% CaCO Data in table I show that 98 percent of the calcite in the froth product of the kaolin flotation operation was recovered in the float product in the form of a concentrate containing only 0.34 percent TiO The data show that 95.5 percent of the anatase in the froth product of the kaolin operation reported in the flotation tailing, producing a concentrate of 62.ll percent TiO grade.

These data therefore show that by aging the froth concentrate from kaolin flotation with a concentrated dispersant solution at high solids, an excellent separation of the calcite from the anatase was effected by flotation without addition of oiling reagents or solvents. The results also show that anatase was selectively depressed by the high solids aging in the presence of the deflocculating agents.

EXAMPLE Il Example I was repeated with the principal exception that the filter cake obtained after washing clay from the froth product of the commercial kaolin flotation contained 73 per cent solids. (In Example I the cake was at 65 percent solids).) After dispersing the cake with 13 lb./ton sodium carbonate and 26 lb./ton N" sodium silicate, the dispersant reagent concentration was 72.2 g./l., in contrast to example I in which the concentration was only 55.7 g./l.

Using the higher solids with more concentrated reagents and employing a froth assaying 8.34 percent TiO the calcite recovery was 98.1 percent. The calcite concentrate analyzed 0.07 percent TiO, and was therefore much purer than it was with the more dilute pulp. The brightness of the calcite concentrate was 88.3 percent as compared to fresh calcite which had a brightness of 91 .0 percent. Using the higher solids pulp, the tailing assayed 77.52 percent TiO- representing 99.2 percent of the 'liO content ofthe froth from the kaolin flotation cells. In contrast. with the more dilute pulp in example l, the

tailing analyzed 66.14 percent TiO at a similar weight recovery.

These results indicate that the anatase was more effectively depressed at the higher solids and that a sharper separation of calcite and anatase was effected when the pulp was aged with dispersant at 73 percent solids than when the pulp was aged with the same dispersants at 65 percent solids.

EXAMPLE lll Employing an Ultraflotation froth product substantially as described in example I, and filtering the froth to 73 to 76 percent solids, various deflocculating agents were added to the filter cakes and attempts were made to float the calcite from the anatase and clay by the procedure employed in examples I and II. Employing 1.61 grams sodium hydrosulfite or l.6l grams sodium sulfide per 200 grams filter cake solids, aging for 18 hours, diluting to 5 percent solids and aerating, there was substantial selective flotation of calcite from the anatase and clay. Employing sodium carbonate as the sole dispersant in amount of 13 lb./ton, there was some selective flotation of calcite but the results were distinctly inferior to the results obtained with 25 lb./ton N sodium silicate alone or a mixture of 26 lb./ton N and i3 lb./ton sodium carbonate. Using 13 lb./ton tetrasodium pyrophosphate or 13 lb./ton sodium hexametaphosphate, separation was effected although the results were inferior to those obtained with the sodium silicate.

EXAMPLE IV A water-washed froth product from the commercial clay Ultraflotation plant was filtered to 75 percent solids. To 200 grams of the filtered froth, 1.8 grams of lithium carbonate was added and the mixture agitated. The mixture, which has the appearance of a thin milk shake, was covered and allowed to age at room temperature for 48 hours. Water was then added in amount sufficient to dilute the aged pulp to 5 percent solids and the diluted pulp was agitated without aeration for 3 minutes. The pulp was floated three times in a 1,000 gram Air- Float flotation cell, producing a float product consisting of substantially pure calcite and representing substantially all of the calcite in the froth product from the Ultraflotation plant. The combined machine discharge product or tailings contained clay and substantially all of the anatase in the froth product from the clay flotation plant.

EXAMPLE V Gray Georgia kaolin (a very fine particle size of sedimentary hard clay having a distinct gray tinge) was subjected to Ultraflotation concentration substantially as described in example I. An Ultraflotation froth concentrate composed of an oiled mixture of Drikalite and anatase was obtained. This froth concentrate was filtered. The filter cake (80 percent solids) was deflocculated by adding dry sodium carbonate l3 lb./ton), mixing, aging for 2 hours at ambient temperature, and then adding N" sodium silicate (26 lb./ton) and aging for IS hours.

The aged pulp was diluted to about 5 percent solids, agitated in a Fagergren cell without aeration for 5 minutes and then floated by introducing air in the pulp. The flotation treatment was very effective in bringing about a sharp separation since with only two cleanings an 87.0 percent brightness calcite product was obtained as the float product. The tailings was a Titania concentrate substantially free from calcite.

EXAMPLE Vl a. A sample of Georgia kaolin crude weighing 6,890 grams, corresponding to 6,000 dry clay, was blunged with 8,000 ml. deionized water, producing a 40 percent solids pulp. After being agitated for 10 minutes, the pulp was dispersed by adding 240 ml. of a 5 percent solution of 0" sodium silicate, corresponding to 4 lb./ton clay. The pulp was then agitated for 10 minutes. To the pulp, 240 ml. of a 5 percent solution of sodium carbonate was added, corresponding to 4 lb./ton clay. The pulp was agitated for 10 minutes and degritted by permitting the pulp to stand for 5 minutes and decanting the supernatant slip from the gritty sediment. The supernatant slip was then fractionated on a centrifuge to produce a slip calculated to contain at least 80 percent by weight of particles finer than 2 microns. The pH of the slip was 9.0. In a control flotation test, 2,280 grams of the slip of fractionated clay at 20.9 percent solids (500 gram dry clay) was conditioned for flotation by adding the following: 100 grams Drikalite; 30 ml. of a 5 percent aqueous solution of ammonium sulfate (corresponding to 6.0 lb./ton clay); 30 ml. of a 25 percent solution of ammonium hydroxide (corresponding to 3.0 lb./ton) and 90 drops of a 50-50 mixture of refined tall oil (M-28") and mineral oil solution of calcium salt of petroleum sulfonate Neutral Calcium Petronate"). The quantity of mixture corresponds to 9.0 lb./ton clay. Seventy-one drops of petroleum hydrocarbon oil (Eureka M") corresponding to 8.0 1b./ton, was added. The pulp was conditioned for 20 minutes and had a pH of 8.9.

The conditioned pulp was floated in a Fagergren flotation cell, removing a froth for minutes. The froth was refloated three times and the machine discharge products combined.

The brightness of the fractionated feed, flotation beneficiated fractionated feed and bleached (zinc hydrosulfite) flotation beneficiated clay were measured. A comparison of the results shows that the brightness of the clay was increased from 81.8 to 88.4 percent by the flotation treatment and further brightened to 90.1 percent by the reducing bleach treatment. The machine discharge contained 452 grams dry clay, representing a 90.4 percent weight recovery.

b. Using the recovered, dried calcite from example II (88.3 percent brightness) in place of the fresh Drikalite, the procedure of part (a) of this example was duplicated. After conditioning the pH of the pulp was 9.0. The recovery and brightness of the flotation beneficiated clay that was conditioned with the reused carrier agent were substantially the same as the recovery and brightness using the fresh calcite carrier, thereby demonstrating that recovered calcite was a suitable substitute for the fresh calcite.

EXAMPLE VI] A gray Georgia kaolin (80.9 percent G.E. brightness) was subjected to Ultraflotation concentration, substantially as described in example VI, substituting 75 grams powdered polyvinyl chloride for the *Drikalite carrier used in example VI. The plastic carrier was composed of particles within the range of I to 3 microns. Ninety percent of the clay was recovered. The beneficiated clay had an unbleached brightness of 85.5 percent and a bleached brightness of 91.1 percent.

a. The forth product of the Ultraflotation concentration (calculated to contain 84.1 grams solids) was allowed to stand, resulting in the suspension of the solids in the froth concentrate. The liquid was siphoned off and the froth solids washed by diluting them with water to 5 percent solids and siphoning the water. This was repeated three times in order to remove clay. The froth was then filtered to 75 percent solids. To the filter cake, dry soda ash was added in amount corresponding to 13 lb./ton of solids in the cake, followed by agitation and the addition of 26 lb./ton of 0" sodium silicate. The filter cake, which was fluidized by addition of the dispersant, was allowed to stand for 18 hours. The aged dispersion was mixed, aerated and floated three times without addition of reagents. The flotation tailings were combined. Analyses of the flotation products showed that 85 percent of the plastic carrier was recovered in the float product. The flotation tailings included the colored impurities originally in the clay.

b. Similar results were obtained when the Ultraflotation froth product was aged in the sodium silicate-sodium carbonate solution at 75 percent solids for 2 hours at 180 F.

Thus, it has been shown that oiled carrier flotation reagents can be selectively floated froinan oiled slimed metal oxide addition, it has been demonstrated that the reclaimed carrier can be recycled and be used as the carrier in a subsequent Ultraflotation beneficiation operation.

Examples VIII and IX illustrate the application of the process of the invention to the beneficiation of various slimed ores in which the minerals separated by the dispersant treatment of a bulk float mineral were present in the ore per se. In Example VIII the bulk float contained an oxide of transition metal element (pyrolusite) and calcite. As mentioned hereinabove, this specific embodiment of the invention is claimed in our copending application Ser. No. 754,95 I. In Example IX, the bulk float was composed largely of sulfide minerals, principally zinc sulfide and tin sulfide; these minerals were separated from each other and recovered in accordance with principles of this invention. This specific embodiment is claimed in US. Pat. No. 3,454,161.

For purposes of comparison, results of conventional flotation beneficiation processes are also given in examples VIII and IX.

EXAMPLE VIII A. Process of the Invention In accordance with this invention a concentrate assaying 52.4 percent Mn was obtained at an overall recovery of 86 percent from a low-grade finely mineralized manganese ore from a deposit in the district of Corral Quemado, Chile. The ore assayed 23.1 percent Mn of which more than 90 percent was present as pyrolusite. Small amounts of manganese silicates were also present. Gangue was predominantly calcite, quartz and silicates. The ore also contained barium sulfate and various carbonate minerals.

The manganese ore was crushed to minus 8 mesh and wetground in a pebble mill at 50 percent solids to 98 percent minus 200 mesh. To remove soluble salts, the ground ore was diluted to 10 percent solids with water and the diluted pulp was allowed to settle. Supernatural liquid was removed by decantation, leaving a washed pulp at about 25 percent solids.

The pulp was dispersed by adding solid sodium carbonate in amount of 1.0 lb./ton and then a hydrosol obtained by diluting 0 sodium silicate solution to 5 percent adding a 1 percent solution of alum The hydrosol was employed in amount equivalent to 8.0 lb./ton 0" sodium silicate and 0.8 lb./ton alum. After addition of each reagent, the pulp was thoroughly agitated.

After the pulp had been dispersed, ammonium sulfate was added in amount of 9 lb./ton. An alkaline collector emulsion was added, following which the pulp was conditioned for 5 minutes with a high-energy input in a Denver Sub A flotation cell. The emulsion was prepared in a Waring Blendor by mixing water with the equivalent of 2.0 lb./ton ammonium hydroxide, 4.5 lb./ton of crude tall oil acids containing about 75 percent fatty acid and 25 percent resin acids and 4.5 lb./ton Calcium Petronate." The emulsion contained about percent water. After the emulsion had been added, fuel oil (Eureka M) was added to the reagentized pulp in amount of 8.0 Ib./ton. The pulp was then conditioned for 20 minutes.

A bulk float of manganese oxide and calcite gangue was obtained by aerating the conditioned pulp in a Denver Sub A flotation cell. After withdrawing a froth for 7 minutes, the float product was cleaned twice by reflotation without addition of reagents. In each cleaner flotation, the pulp was diluted to maintain adequate pulp level in the flotation cell. The flotation tailings from the bulk float operation were discarded and the final float product (the concentrate of pyrolusite and calcite) was dewatered by filtration, resulting in a filter cake containing about 70 percent solid. I

The filter cake was charged to a pug mill and deflocculated and thinned by adding dry sodium carbonate in amount corresponding to 13 lb./ton, followed by addition of 26 lb./ton sodium silicate. The pug mill was in operation while the dispersing agents were added. The resulting creamy mass was held in a container at room temperature for 18 hours without agitation.

To separate the manganese oxide from the carbonate minerals in the mass, the dispersant-treated bulk float was diluted to about l0 percent solids and dextrine was added to help depress the manganese. After conditioning the pulp for minutes in a Denver Sub A flotation cell, air was admitted and a froth was withdrawn for 5 minutes. The froth was cleaned twice by flotation and the three machine discharge products were combined to produce the manganese concentrate.

An X-ray diffraction pattern of the manganese concentrate indicated that it was composed predominantly of pyrolusite and contained only traces of calcite. The concentrate represented 39.9 percent by weight of the starting ore and analyzed 52.4 percent Mn, corresponding to a pyrolusite product of about 85 percent purity. Overall recovery of manganese was 86.0 percent.

B. Prior Art Treatment To illustrate the advantages of the process of the invention over prior art processes for concentrating manganese, the metallurgical results obtained in part A. of this example (process of the invention) were compared to results obtained when samples of ore from the same deposit were concentrated by prior art flotation methods.

One of the prior procedures involved grinding the ore to 70 percent minus 200 mesh, followed by flotation of calcite from manganese using an oleic acid collector, fuel oil and quebracho. With an ore containing 23.4 percent Mn, a concentrate analyzing 42.9 percent Mn (69.2 percent purity) was obtained at a recovery of 69.8 percent. With a similar sample of the ore ground to 90 percent minus 200 mesh and flotation of calcite from the manganese with a fatty acid collector, dextrine to depress manganese and pine oil frother, the concentrate analyzed 43.5 percent Mn and represented a 67.7 percent recovery.

A comparison of these results with the metallurgical results for the process of the invention which appear in part A. of this example, shows that about 20 percent more manganese was recovered by the process of the invention and that the manganese was recovered as a concentrate containing about 9 percent more manganese. Thus, the process of the invention was superior to the prior art processes with respect to both the grade and recovery of manganese.

EXAMPLE [X This example illustrates separations of zinc and tin from a complex zinc-tin Bolivian sulfide ore containing sphalerite, stannite (present as inclusions in the sphalerite), teallite (a solid solution of the composition PbS.SnS and tuffahlite (a zinc sulfide-tin sulfide mineral). Small amounts of cassiterite were also present. Gangue minerals include galena, pyrite, quartz and aluminosilicates.

A petrographic inspection of a representative sample of the ore indicated that the zinc and tin sulfide minerals were present in a state of extremely fine dissemination.

Chemical assays of a representative sample showed the ore analyzed l4.37% by weight Zn, 2.03% Sn, 2.01% Pb, l4.54% Fe O 7.62% M 0 41.25% total SiO and 20.06% free SiO A. Process of the Invention A minus 8 mesh sample of the Bolivian zinc-tin ore was ground at 50 percent solids in a pebble mill to I00 percent minus 325 mesh. The ground ore pulp was diluted with water to l0 percent solids and the diluted pulp was flocculated by adding sulfuric acid to a pH of 3.5. Supernatant liquid was decanted and the thickened pulp was fluidized and dispersed by adding 10 lb./ton soda ash and 3 lb./ton O sodium silicate. The dispersed pulp was conditioned for flotation by adding 6.0

lb./ton ammonium sulfate as a 5 percent aqueous solution and an emulsion containing water, l.0 lb./ton ammonium hydroxide, 6.2 lb./ton crude tall oil and 6.2 lb./ton Calcium Petronate. The pulp was conditioned for 20 minutes and floated in a subaeration-type flotation cell.

The bulk cleaner froth concentrate was cleaned three times by reflotation.

The tailings were discarded and the cleaned bulk concentrate was filtered. The filter cake which contained about 70 percent solids was fluidized by incorporating l3 lb./ton soda ash and 26 lb./ton 0" sodium silicate solution. The fluidized cake was maintained in a closed container for 20 hours.

The aged, dispersed filter cake was then diluted to about l5 percent solids and aerated in the subaeration flotation machine. A froth product (the zinc concentrate) was withdrawn and recleaned twice by flotation without addition of reagents. The three tailings were combined, forming the tin concentrate.

A summary of the overall results appears in table II.

TABLE ll PROCESS OF THE lNVENTlON-FLOTATION Data in table II show that 68.5 percent of the tin in the ore was recovered in the final tin concentrate containing 5.03 percent Sn. The recovery of zinc in the zinc concentrate was about 65.9 percent and this concentrate contained 45.68 percent Zn by weight.

B. Prior Art Treatment A sample of the complex sulfide ore was crushed to minus 8 mesh (Tyler) and ground in a stainless steel rod mill at 50 percent solids in the presence of soda ash (l5 lb./ton) to prevent activation of sphalerite and pyrite by iron from the mill. During the g inding, 0.6 lb./ton sodium cyanide and 2.0 lb./ton zinc sulfate were added to promote the deactivation of sphalerite. After grinding, the ore was diluted with water to 20 percent solids. The pulp was conditioned for flotation of stannite by adding sodium hydroxide to a pH of 9.5 and 0.025 lb./ton Z-l l" xanthate. A rougher tin-lead flotation was made. The rougher float was cleaned three times without addition of reagents, producing the final tin-lead concentrate and a combined cleaner tails. The rougher tailings were treated with 1.5 lb./ton copper sulfate pentahydrate to reactivate the sphalerite and 3.0 lb./ton lime for pH control. The pulp was then conditioned for sphalerite flotation with 0.075 lb./ton Z-l l" xanthate. A small amount of Dowfrother 250 was added. A second float was taken after addition of 0.075 lb./ton Z-l l" xanthate. The float products were combined to make the zinc rougher concentrate. The flotation tailings were discarded. The zinc rougher concentrate was treated with 2.4 lb./ton of lime and was cleaned twice, producing a zinc cleaner concentrate and a zinc cleaner tails product.

Metallurgical results for the test are summarized in table lll.

TABLE III Zinc Conc.

Tin Cone.

The data in table lll show that 75.1 percent of the Zn was recovered in the form of a concentrate of 54.22 percent Zn grade by the conventional sulfide process. However, only 5.5 percent of the Sn values were recovered in the form of a concentrate of 3.65 percent Sn grade. The results therefore indicate that the zinc mineral responded well to the conventional sulfide flotation but that tin did not.

A comparison of the data in table ll (process of the invention) with data in table lll (conventional sulfide flotation with depressant for sphalerite) shows that with the process of the invention the recovery of tin in the tailings was 68.5 percent, while in the conventional tin flotation process only 5.5 percent of the tin was recovered in the tin concentrate. The data show also that the tin grade was higher when the separation was made in accordance with the process of the invention. Zinc recovery and grade were slightly lower with the process of the invention. Thus, the process of the invention resulted in a marked improvement in tin recovery and grade without substantial sacrifice in zinc recovery or grade.

We claim:

1. In the beneficiation of an ore wherein an aqueous ore pulp containing finely divided particles of alkaline earth carbonate mineral and finely divided particles of a mineral which is an oxide of a transition metal is subjected to flotation in the presence of a fatty acid collector reagent selective to said alkaline earth carbonate mineral and said oxide of a transition metal, thereby forming a bulk float product which is a fatty acid reagentized mixture of said minerals, an improved method for separating said reagentized minerals in said bulk float product from each other which comprises:

removing sufficient water from said bulk float product to form a mass having a solids content of at least 50 percent by weight, incorporating an alkaline deflocculating agent with said mass in amount sufficient to thin it while restricting the amount of water to an amount such that the solids content of said mass is at least 50 percent by weight, aging said mass at ambient temperature for at least l2 hours, diluting the aged mass with water to form a pulp, aerating the diluted pulp, thereby forming a froth product which is a concentrate of alkaline earth carbonate mineral, and separating said froth from a tailing which is a concentrate of said transition metal oxide mineral.

2. The method of claim 1 in which said alkaline earth carbonate is calcium carbonate.

3. The method of claim 2 wherein said mass has a solids content within the range of 70 to percent during said aging.

4. The method of claim 2 wherein both of said minerals contain a substantial portion of minus 325 mesh particles.

5. The method of claim 4 wherein said oxide mineral is pyrolusite.

6. The method of claim 4 wherein said alkaline deflocculating agent comprises sodium silicate and wherein total concentration of deflocculating agent is at least 50 grams/liter.

7. The method of claim 6 wherein said carbonate mineral includes calcium carbonate present as an ore constituent.

8. The method of claim 4 wherein said oxide mineral is Titania.

9. The method of claim 8 wherein said Titania is present in the form of yellow colored anatase.

10. The method of claim 9 wherein said alkaline earth metal carbonate mineral is calcite which is added to an ore pulp containing said anatase for the purpose of aiding the flotation of said anatase with said fatty acid collector.

Patent No. 3,635,337 Dated January 2 Inv'entofls) Venancio Mercade and Samuel R. Weir It is certified that error appears in the above-identified patent and that said Letters Patent are hereby corrected as shown below:

Column 1 page 1 line [72] "Venacio Mercade" should read Venancio Mercade 7 Column 5 line 26, "0" sodium silicate solution" should read "O" sodium silicate solution Column 6 line 46, "Petronate" should read Petronate line 71, "Q" sodium silicate solution" should read "0" sodium silicate solution Column 7 line 62, "N" sodium silicate" should read "N" sodium silicate Column 8 line 70, "6,000 dry clay" should read 6,000 grams dry clay Column 12 Table III, add the following column headings Wt. Z Sn 2 Zn Z Sn Zn Zinc conc. 20.23 0.85 54.22 9.2 I 75.1

Tin conc. 2.80 3.65 16.38 5.5 3.2

Signed and sealed this 27th day of June 1972.

(SEAL) Attest:

EDWARD M.FLETCHER,JR. ROBERT GOTTSCHALK Attssting Officer Commissioner of Patents FORM PO-105O (10-69) USCQMM-DC 50375-p5 w u.s. GOVERNMENT PRINTING ornc: was 0-365-334

Claims (9)

1. 2. The method of claim 1 in which said alkaline earth carbonate is calcium carbonate.
2. 3. The method of claim 2 wherein said mass has a solids content within the range of 70 to 80 percent during said aging.
3. 4. The method of claim 2 wherein both of said minerals contain a substantial portion of minus 325 mesh particles.
4. 5. The method of claim 4 wherein said oxide mineral is pyrolusite.
5. 6. The method of claim 4 wherein said alkaline deflocculating agent comprises sodium silicate and wherein total concentration of deflocculating agent is at least 50 grams/liter.
6. 7. The method of claim 6 wherein said carbonate mineral includes calcium carbonate present as an ore constituent.
7. 8. The method of claim 4 wherein said oxide mineral is Titania.
8. 9. The method of claim 8 wherein said Titania is present in the form of yellow colored anatase.
9. 10. The method of claim 9 wherein said alkaline earth metal carbonate mineral is calcite which is added to an ore pulp containing said anatase for the purpose of aiding the flotation of said anatase with said fatty acid collector.