Alkylated diaryl oxide monosulfonate collectors useful in the flotation of minerals

Abstract: Alkylated diaryl oxide monosulfonic acids or salts thereof or their mixture are useful as collectors in the flotation of minerals, particularly oxide minerals.

Description: This invention is related to the recovery of minerals by froth flotation. Flotation is a process of treating a mixture of finely divided mineral solids, e.g., a pulverulent ore, suspended in a liquid whereby a portion of the solids is separated from other finely divided mineral solids, e.g., silica, siliceous gangue, clays and other like materials present in the ore, by introducing a gas (or providing a gas in situ ) in the liquid to produce a frothy mass containing certain of the solids on the top of the liquid, and leaving suspended (unfrothed) other solid components of the ore. Flotation is based on the principle that introducing a gas into a liquid containing solid particles of different materials suspended therein causes adherence of some gas to certain suspended solids and not to others and makes the particles having the gas thus adhered thereto lighter than the liquid. Accordingly, these particles rise to the top of the liquid to form a froth. The minerals and their associated gangue which are treated by froth flotation generally do not possess sufficient hydrophobicity or hydrophilicity to allow adequate separation. Therefore, various chemical reagents are often employed in froth flotation to create or enhance the properties necessary to allow separation. Collectors are used to enhance the hydrophobicity and thus the floatability of different mineral values. Collectors must have the ability to (1) attach to the desired mineral species to the relative exclusion of other species present; (2) maintain the attachment in the turbulence or shear associated with froth flotation; and (3) render the desired mineral species sufficiently hydrophobic to permit the required degree of separation. A number of other chemical reagents are used in addition to collectors. Examples of types of additional reagents used include frothers, depressants, pH regulators, such as lime and soda, dispersants and various promoters and activators. Depressants are used to increase or enhance the hydrophilicity of various mineral species and thus depress their flotation. Frothers are reagents added to flotation systems to promote the creation of a semi-stable froth. Unlike both depressants and collectors, frothers need not attach or adsorb on mineral particles. Froth flotation has been extensively practiced in the mining industry since at least the early twentieth century. A wide variety of compounds are taught to be useful as collectors, frothers and other reagents in froth flotation. For example, xanthates, simple alkylamines, alkyl sulfates, alkyl sulfonates, carboxylic acids and fatty acids are generally accepted as useful collectors. Reagents useful as frothers include lower molecular weight alcohols such as methyl isobutyl carbinol and glycol ethers. The specific additives used in a particular flotation operation are selected according to the nature of the ore, the conditions under which the flotation will take place, the mineral sought to be recovered and the other additives which are to be used in combination therewith. While a wide variety of chemical reagents are recognized by those skilled in the art as having utility in froth flotation, it is also recognized that the effectiveness of known reagents varies greatly depending on the particular ore or ores being subjected to flotation as well as the flotation conditions. It is further recognized that selectivity or the ability to selectively float the desired species to the exclusion of undesired species is a particular problem. Minerals and their associated ores are generally categorized as sulfides or oxides, with the latter group comprising oxygen-containing species such as carbonates, hydroxides, sulfates and silicates. Thus, the group of minerals categorized as oxides generally include any oxygen-containing mineral. While a large proportion of the minerals existing today are contained in oxide ores, the bulk of successful froth flotation systems is directed to sulfide ores. The flotation of oxide minerals is recognized as being substantially more difficult than the flotation of sulfide minerals and the effectiveness of most flotation processes in the recovery of oxide ores is limited. A major problem associated with the recovery of both oxide and sulfide minerals is selectivity. Some of the recognized collectors such as the carboxylic acids, alkyl sulfates and alkyl sulfonates discussed above are taught to be effective collectors for oxide mineral ores. However, while the use of these collectors can result in acceptable recoveries, it is recognized that the selectivity to the desired mineral value is typically quite poor. That is, the grade or the percentage of the desired component contained in the recovered mineral is unacceptably low. Due to the low grade of oxide mineral recovery obtained using conventional, direct flotation, the mining industry has generally turned to more complicated methods in an attempt to obtain acceptable recovery of acceptable grade minerals. Oxide ores are often subjected to a sulfidization step prior to conventional flotation in existing commercial processes. After the oxide minerals are sulfidized, they are then subjected to flotation using known sulfide collectors. Even with the sulfidization step, recoveries and grade are less than desirable. An alternate approach to the recovery of oxide ores is liquid/liquid extraction. A third approach used in the recovery of oxide ores, particularly iron oxides and phosphates, is reverse or indirect flotation. In reverse flotation, the flotation of the ore having the desired mineral values is depressed and the gangue or other contaminant is floated. In some cases, the contaminant is a mineral which may have value. A fourth approach to mineral recovery involves chemical dissolution or leaching. US-A-4172029 discloses the use of sulforated polycyclic compounds in froth flotation of phosphate ores. None of these existing methods of flotation directed to oxide ores are without problems. Generally, known methods result in low recovery or low grade or both. The low grade of the minerals recovered is recognized as a particular problem in oxide mineral flotation. Known recovery methods have not been economically feasible and consequently, a large proportion of oxide ores simply are not processed. Thus, a great need for improved selectivity in oxide mineral flotation is generally acknowledged by those skilled in the art of froth flotation. The present invention is a process for the recovery of minerals by froth flotation comprising subjecting an aqueous slurry comprising particulate minerals to froth flotation in the presence of a collector comprising an alkylated a diaryl oxide sulfonic acid, preferably an alkylated diphenyl oxide sulfonic acid or a salt thereof, or a mixture of such salts or acids, wherein monosulfonated species comprise at least about 20 weight percent of the sulfonated acids or salts, under conditions such that the minerals to be recovered are floated. The recovered minerals can be the mineral that is desired or can be undesired contaminants. Additionally, the froth flotation process of this invention may utilize frothers and other flotation reagents known in the art. The practice of the flotation process of this invention results in improvements in selectivity and thus the grade of minerals recovered from oxide and/or sulfide ores while generally maintaining or increasing overall recovery levels of the desired mineral. It is surprising that the use of an alkylated diphenyl oxide sulfonic acid such as alkylated diphenyl oxide monosulfonic acids or salts thereof results in consistent improvements in selectivity or recovery of mineral values. The flotation process of this invention is useful in the recovery of mineral values from a variety of ores, including oxide ores as well as sulfide ores and mixed ores. Non-limiting examples of oxide ores which can be floated using the practice of this invention preferably include iron oxides, nickel oxides, copper oxides, phosphorus oxides, aluminum oxides and titanium oxides. Other types of oxygen-containing minerals which may be floated using the practice of this invention include carbonates such as calcite or dolomite and hydroxides such as bauxite. Non-limiting examples of specific oxide ores which can be collected by froth flotation using the process of this invention include those containing cassiterite, hematite, cuprite, vallerite, calcite, talc, kaolin, apatite, dolomite, bauxite, spinel, corundum, laterite, azurite, rutile, magnetite, columbite, ilmenite, smithsonite, anglesite, scheelite, chromite, cerussite, pyrolusite, malachite, chrysocolla, zincite, massicot, bixbyite, anatase, brookite, tungstite, uraninite, gummite, brucite, manganite, psilomelane, goethite, limonite, chrysoberyl, microlite, tantalite, topaz and samarskite. The froth flotation process of this invention will be useful for the processing of additional ores including oxide ores, wherein oxide is defined to include carbonates, hydroxides, sulfates and silicates as well as oxides. The process of this invention is also useful in the flotation of sulfide ores. Non-limiting examples of sulfide ores which can be floated by the process of this invention include those containing chalcopyrite, chalcocite, galena, pyrite, sphalerite, molybdenite and pentlandite. Noble metals such as gold and silver and the platinum group metals wherein platinum group metals comprise platinum, ruthenium, rhodium, palladium, osmium, and iridium, can also be recovered by the practice of this invention. For example, such metals are sometimes found associated with oxide and/or sulfide ores. Platinum, for example, can be found associated with troilite. By the practice of the present invention, such metals can be recovered in good yield. Ores do not always exist purely as oxide ores or as sulfide ores. Ores occurring in nature can comprise both sulfur-containing and oxygen-containing minerals as well as small amounts of noble metals as discussed above. Minerals can be recovered from these mixed ores by the practice of this invention. This can be done in a two-stage flotation where one stage comprises conventional sulfide flotation to recover primarily sulfide minerals and the other stage of the flotation utilizes the process and collector composition of the present invention to recover primarily oxide minerals and any noble metals that may be present. Alternatively, both the sulfur-containing and oxygen-containing minerals can be recovered simultaneously by the practice of this invention. A particular feature of the process of this invention is the ability to differentially float various minerals. Without wishing to be bound by theory, it is thought that the susceptibility of various minerals to flotation in the process of this invention is related to the crystal structure of the minerals. More specifically, a correlation appears to exist between the ratio of crystal edge lengths to crystal surface area on a unit area basis. Minerals having higher ratios appear to float preferentially when compared to minerals having lower ratios. Thus, minerals whose crystal structure has 24 or more faces (Group I) are generally more likely to float than minerals having 16 to 24 faces (Group II). Group III minerals comprising minerals having 12 to 16 faces are next in order of preferentially floating followed by Group IV minerals having 8 to 12 faces. In the process of this invention, generally Group I minerals will float before Group II minerals, which will float before Group III minerals, which will float before Group IV minerals. By floating before or preferentially floating, it is meant that the preferred species will float at lower dosages (amount) of collector that is used. That is, a Group I mineral can be collected at a very low dosage. Upon increasing the dosage and/or the removal of most of the Group I mineral, a Group II mineral will be collected and so on. One skilled in the art will recognize that these groupings are not absolute. Various minerals can have different possible crystal structures. Further the size of crystals existing in nature also varies which will influence the ease with which different minerals can be floated. An additional factor affecting flotation preference is the degree of liberation. Further, within a group, that is, among minerals whose crystals have similar edge length to surface area ratios, these factors and others will influence which member of the group floats first. One skilled in the art can readily determine which group a mineral belongs to by examining standard mineralogy characterization of different minerals. These are available, for example, in Manual of Mineralogy, 19th Edition , Cornelius S. Hurlbut, Jr. and Cornelis Klein (John Wiley and Sons, New York 1977). Non-limiting examples of minerals in Group I include graphite, niccolite, covellite, molybdenite and beryl. Non-limiting examples of minerals in Group II include rutile, pyrolusite, cassiterite, anatase, calomel, torbernite, autunite, marialite, meionite, apophyllite, zircon and xenotime. Non-limiting examples of minerals in Group III include arsenic, greenockite, millerite, zincite, corundum, hematite, brucite, calcite, magnesite, siderite, rhodochrosite, smithsonite, soda niter, apatite, pyromorphite, mimetite and vanadinite. Non-limiting examples of minerals in Group IV include sulfur, chalcocite, chalcopyrite, stibnite, bismuthinite, loellingite, marcasite, massicot, brookite, boehmite, diaspore, goethite, samarskite, atacamite, aragonite, witherite, strontianite, cerussite, phosgenite, niter, thenardite, barite, celestite, anglesite, anhydrite, epsomite, antlerite, caledonite, triphylite, lithiophilite, heterosite, purpurite, variscite, strengite, chrysoberyl, scorodite, descloizite, mottramite, brazilianite, olivenite, libethenite, adamite, phosphuranylite, childrenite, eosphorite, scheelite, powellite, wulfenite, topaz, columbite and tantalite. As discussed above, these groupings are theorized to be useful in identifying which minerals will be preferentially floated. However, the collector and process of this invention are useful in the flotation of various minerals which do not fit into the above categories. These groupings are useful in predicting which minerals will float at the lowest relative collector dosage, not in determining which minerals can be collected by flotation in the process of this invention. The selectivity demonstrated by the collectors of this invention permit the separation of small amounts of undesired minerals from the desired minerals. For example, the presence of apatite is frequently a problem in the flotation of iron as is the presence of topaz in the flotation of cassiterite. Thus, the collectors of the present invention are, in some cases, useful in reverse flotation where the undesired mineral is floated such as floating topaz away from cassiterite or apatite from iron. In addition to the flotation of ores found in nature, the flotation process and collector composition of this invention are useful in the flotation of minerals from other sources. One such example is the waste materials from various processes such as heavy media separation, magnetic separation, metal working and petroleum processing. These waste materials often contain minerals that can be recovered using the flotation process of the present invention. Another example is the recovery of a mixture of graphite ink and other carbon based inks in the recycling of paper. Typically such recycled papers are de-inked to separate the inks from the paper fibers by a flotation process. The flotation process of the present invention is particularly effective in such de-inking flotation processes. The alkylated diaryl oxide monosulfonic acid or monosulfonate collector employed in the process of this invention corresponds to the general formula: Ar′-O-Ar wherein Ar′ and Ar are independently in each occurrence substituted or unsubstituted aromatic moieties such as, for example, phenyl or naphthyl with the proviso that one and only one of Ar′ and Ar contain one sulfonic acid or sulfonic acid salt moiety and at least one of Ar′ and Ar is alkylated. Preferably, the diaryl oxide monosulfonic acid or monosulfonate collector is an alkylated diphenyl oxide or an alkylated biphenyl phenyl oxide monosulfonic acid or monosulfonate or mixture thereof. The diaryl oxide monosulfonic acid or monosulfonate is substituted with one or more alkyl substituents. The alkyl substituents can be substituted or unsubstituted alkyl or substituted or unsubstituted unsaturated alkyl. The monosulfonated alkylated diaryl oxide collector of this invention is more preferably a diphenyl oxide collector and corresponds to the following formula or to a mixture of compounds corresponding to the formula: wherein each R is independently a saturated alkyl or substituted saturated alkyl radical or an unsaturated alkyl or substituted unsaturated alkyl radical; each m and n is independently 0, 1 or 2 but at least one of m and n is not 0; each M is independently hydrogen, an alkali metal, alkaline earth metal, or ammonium or substituted ammonium and each x and y are individually 0 or 1 with the proviso that the sum of x and y is one. Preferably, the R group(s) is independently an alkyl group having from 1 to 24, more preferably from 6 to 24 carbon atoms, even more preferably from 6 to 16 carbon atoms and most preferably from 10 to 16 carbon atoms. The alkyl groups can be linear, branched or cyclic with linear or branched radicals being preferred. It is also preferred that m and n are each one. The M⁺ ammonium ion radicals are of the formula (R′)₃HN⁺ wherein each R′ is independently hydrogen, a C₁-C₄ alkyl or a C₁-C₄ hydroxyalkyl radical. Illustrative C₁-C₄ alkyl and hydroxyalkyl radicals include methyl, ethyl, propyl, isopropyl, butyl, hydroxymethyl and hydroxyethyl. Typical ammonium ion radicals include ammonium (N⁺H₄), methylammonium (CH₃N⁺H₃), ethylammonium (C₂H₅N⁺H₃), dimethylammonium ((CH₃)₂N⁺H₂), methylethylammonium (CH₃N⁺H₂C₂H₅), trimethylammonium ((CH₃)₃N⁺H), dimethylbutylammonium ((CH₃)₂N⁺HC₄H₉), hydroxyethylammonium (HOCH₂CH₂N⁺H₃) and methylhydroxyethylammonium (CH₃N⁺H₂CH₂CH₂OH). Preferably, each M is hydrogen, sodium, calcium, potassium or ammonium. Alkylated diphenyl oxide sulfonates and their methods of preparation are well-known and reference is made thereto for the purposes of this invention. The monosulfonate collectors of the present invention can be prepared by modifications to known methods of preparation of sulfonates. Representative methods of preparation of sulfonates are disclosed in US-A- 3,264,242; 3,634,272; and 3,945,437. Commercial methods of preparation of the alkylated diphenyl oxide sulfonates generally do not produce species which are exclusively monoalkylated, monosulfonated, dialkylated or disulfonated. The commercially available species are predominantly (greater than 90 percent) disulfonated and are a mixture of mono- and dialkylated with the percentage of dialkylation being from 15 to 25 and the percentage of monoalkylation being from 75 to 85 percent. Most typically, the commercially available species are about 80 percent monoalkylated and 20 percent dialkylated. In the practice of this invention, the use of monosulfonated species has been found to be critical. Such monosulfonated species can be prepared by a modification of the sulfonation step in the methods described in, for example, US-A- 3,264,242; 3,634,272; and 3,945,437. Specifically, the methods taught above are directed to preparing predominantly disulfonated species. Thus, in the sulfonation step, it is taught to use sufficient sulfonating agent to sulfonate both aromatic rings. However, in the preparation of the monosulfonates useful in the practice of the present invention, the amount of sulfonating agent used is preferably limited to that needed to provide one sulfonate group per molecule. The monosulfonates prepared in this way will include both molecules which are not sulfonated as well as those which contain more than one sulfonate group per molecule. If desired, the monosulfonates can be separated and used in relatively pure form. However, the mixture resulting from a sulfonation step utilizing only sufficient sulfonating agent to provide approximately one sulfonate group per molecule is also useful in the practice of this invention. As stated above, the use of monosulfonated species is critical to the practice of this invention. However, the presence of disulfonated species is not thought to be detrimental from a theoretical standpoint as long as at least 20 weight percent of the monosulfonated species is present. It is preferred that at least 25 weight percent monosulfonation is present and more preferred that at least 40 weight percent monosulfonation is present and most preferred that at least 50 weight percent monosulfonation is present. It is most preferred to use relatively pure monosulfonated acids or salts. In commercial applications, one skilled in the art will recognize that whatever higher costs are associated with the production of the relatively pure monosulfonated species will be balanced against decreases in effectiveness associated with the use of mixtures containing disulfonated species. Commercially available alkylated diphenyl oxide sulfonates frequently are mixtures of monoalkylated and dialkylated species. While such mixtures of monoalkylated and dialkylated species are operable in the practice of this invention, it is preferable in some circumstances to use species that are either monoalkylated, dialkylated or trialkylated. Such species are prepared by modifications of the methods described in, for example, US-A- 3,264,242; 3,634,272; and 3,945,437. When it is desired to use other than a mixture, a distillation step is inserted after alkylation to remove monoalkylated species and either use the monoalkylated species or recycle it for further alkylation. Generally, it is preferred to use dialkylated species although monoalkylated and trialkylated are operable. Non-limiting examples of preferred alkylated diphenyl oxide sulfonates include sodium monosulfonated diphenyl oxide, sodium monosulfonated hexyldiphenyl oxide, sodium monosulfonated decyldiphenyl oxide, sodium monosulfonated dodecyldiphenyl oxide, sodium monosulfonated hexadecyldiphenyl oxide, sodium monosulfonated eicosyldiphenyl oxide and mixtures thereof. In a more preferred embodiment, the collector is a sodium monosulfonated dialkylated diphenyl oxide wherein the alkyl group is a C₁₀₋₁₆ alkyl group, most preferably a C₁₀₋₁₂ alkyl group. The alkyl groups can be branched or linear. The collector can be used in any concentration which gives the desired selectivity and recovery of the desired mineral values. In particular, the concentration used is dependent upon the particular mineral to be recovered, the grade of the ore to be subjected to the froth flotation process and the desired quality of the mineral to be recovered. Additional factors to be considered in determining dosage levels include the amount of surface area of the ore to be treated. As will be recognized by one skilled in the art, the smaller the particle size, the greater the surface area of the ore and the greater the amount of collector reagents needed to obtain adequate recoveries and grades. Typically, oxide mineral ores must be ground finer than sulfide ores and thus require very high collector dosages or the removal of the finest particles by desliming. Conventional processes for the flotation of oxide minerals typically require a desliming step to remove the fines present and thus permit the process to function with acceptable collector dosage levels. The collector of the present invention functions at acceptable dosage levels with or without desliming. Preferably, the concentration of the collector is at least 0.001 kg/metric ton, more preferably at least 0.05 kg/metric ton. It is also preferred that the total concentration of the collector is no greater than 5.0 kg/metric ton and more preferred that it is no greater than 2.5 kg/metric ton. In general, to obtain optimum performance from the collector, it is most advantageous to begin at low dosage levels and increase the dosage level until the desired effect is achieved. While the increases in recovery and grade obtained by the practice of this invention increase with increasing dosage, it will be recognized by those skilled in the art that at some point the increase in recovery and grade obtained by higher dosage is offset by the increased cost of the flotation chemicals. It will also be recognized by those skilled in the art that varying collector dosages are required depending on the type of ore and other conditions of flotation. Additionally, the collector dosage required has been found to be related to the amount of mineral to be collected. In those situations where a small amount of a mineral susceptible to flotation using the process of this invention, a very low collector dosage is needed due to the selectivity of the collector. It has been found advantageous in the recovery of certain minerals to add the collector to the flotation system in stages. By staged addition, it is meant that a part of the collector dose is added; froth concentrate is collected; an additional portion of the collector is added; and froth concentrate is again collected. The total amount of collector used is preferably not changed when it is added in stages. This staged addition can be repeated several times to obtain optimum recovery and grade. The number of stages in which the collector is added is limited only by practical and economic constraints. Preferably, no more than about six stages are used. An additional advantage of staged addition is related to the ability of the collector of the present invention to differentially float different minerals at different dosage levels. As discussed above, at low dosage levels, one mineral particularly susceptible to flotation by the collector of this invention is floated while other minerals remain in the slurry. At an increased dosage, a different mineral may be floated thus permitting the separation of different minerals contained in a given ore. In addition to the collector of this invention, other conventional reagents or additives can be used in the flotation process. Examples of such additives include various depressants and dispersants well-known to those skilled in the art. Additionally, the use of hydroxy-containing compounds such as alkanol amines or alkylene glycols has been found to be useful in improving the selectivity to the desired mineral values in systems containing silica or siliceous gangue. The collector of this invention can also be used in conjunction with other collectors. In addition, frothers can be and typically are used. Frothers are well known in the art and reference is made thereto for the purposes of this invention. Examples of useful frothers include polyglycol ethers and lower molecular weight frothing alcohols. A particular advantage of the collector of the present invention is that additional additives are not required to adjust the pH of the flotation slurry. The flotation process utilizing the collector of the present invention operates effectively at typical natural ore pH's ranging from 5 or lower to 9. This is particularly important when considering the cost of reagents needed to adjust slurry pH from a natural pH of around 7.0 or lower to 9.0 or 10.0 or above which is typically necessary using conventional carboxylic, sulfonic, phosphonic and xanthic collectors. The ability of the collector of the present invention to function at relatively low pH means that it can also be used in those instances where it is desired to lower the slurry pH. The lower limit on the slurry pH at which the present invention is operable is that pH at which the surface charge on the mineral species is suitable for attachment by the collector. Since the collector of the present invention functions at different pH levels, it is possible to take advantage of the tendency of different minerals to float at different pH levels. This makes it possible to do one flotation run at one pH to optimize flotation of a particular species. The pH can then be adjusted for a subsequent run to optimize flotation of a different species thus facilitating separation of various minerals found together. The collector of this invention may also be used in conjunction with conventional collectors. For example, the monosulfonated diaryl oxide collectors of this invention may be used in a two-stage flotation in which the monosulfonated diaryl oxide flotation recovers primarily oxide minerals while a second stage flotation using conventional collectors is used to recover primarily sulfide minerals or additional oxide minerals. When used in conjunction with conventional collectors, a two-stage flotation may be used wherein the first stage comprises the process of this invention and is done at the natural pH of the slurry. The second stage involves conventional collectors and is conducted at an elevated pH. It should be noted that in some circumstances, it may be desirable to reverse the stages. Such a two-stage process has the advantages of using less additives to adjust pH and also permits a more complete recovery of the desired minerals by conducting flotation under different conditions. The following examples are provided to illustrate the invention and should not be interpreted as limiting it in any way. Unless stated otherwise, all parts and percentages are by weight. The following examples include work involving Hallimond tube flotation and flotation done in laboratory scale flotation cells. It should be noted that Hallimond tube flotation is a simple way to screen collectors, but does not necessarily predict the success of collectors in actual flotation. Hallimond tube flotation does not involve the shear or agitation present in actual flotation and does not measure the effect of frothers. Thus, while a collector must be effective in a Hallimond tube flotation if it is to be effective in actual flotation, a collector effective in Hallimond tube flotation will not necessarily be effective in actual flotation. It should also be noted that experience has shown that collector dosages required to obtain satisfactory recoveries in a Hallimond tube are often substantially higher than those required in a flotation cell test. Thus, the Hallimond tube work cannot precisely predict dosages that would be required in an actual flotation cell. Example 1 - Hallimond Tube Flotation of Malachite and Silica About 1.1 g of (1) malachite, a copper oxide mineral having the approximate formula Cu₂CO₃(OH)₂, or (2) silica was sized to about -60 to +120 U.S. mesh and placed in a small bottle with about 20 ml of deionized water. The mixture was shaken for 30 seconds and the water phase containing some suspended fine solids or slimes decanted. This desliming step was repeated several times. A 150-ml portion of deionised water was placed in a 250-ml glass beaker. Next, 2.0 ml of a 0.10 molar solution of potassium nitrate was added as a buffer electrolyte. The pH was adjusted to about 10.0 with the addition of 0.10 N HCl and/or 0.10 N NaOH. Next, a 1.0-g portion of the deslimed mineral was added along with deionized water to bring the total volume to about 180 ml. The collector, a branched C₁₆ alkylated sodium diphenyl oxide sulfonate comprising about 80 percent monoalkylated species and about 20 percent dialkylated species, was added and allowed to condition with stirring for 15 minutes. The pH was monitored and adjusted as necessary using HCl and NaOH. It should be noted that Runs 1-5 are not embodiments of the invention and use a disulfonated collector while Runs 6-10, which are embodiments of the invention, use a monosulfonated collector. The only difference in the collectors used in Runs 1-5 and those used in Runs 6-10 is disulfonated versus monosulfonation. The slurry was transferred into a Hallimond tube designed to allow a hollow needle to be fitted at the base of the 180-ml tube. After the addition of the slurry to the Hallimond tube, a vacuum of 5 inches (12.7 cm) of mercury was applied to the opening of the tube for a period of 10 minutes. This vacuum allowed air bubbles to enter the tube through the hollow needle inserted at the base of the tube. During flotation, the slurry was agitated with a magnetic stirrer set at 200 revolutions per minute (RPM). The floated and unfloated material was filtered out of the slurry and oven dried at 100°C. Each portion was weighed and the fractional recoveries of copper and silica reported in Table I below. After each test, all equipment was washed with concentrated HCl and rinsed with 0.10 N NaOH and deionized water before the next run. The recovery of copper and silica, respectively, reported is that fractional portion of the original mineral placed in the Hallimond tube that is recovered. Thus, a recovery of 1.00 indicates that all of the material was recovered. It should be noted that although the recovery of copper and silica, respectively, is reported together, the data is actually collected in two experiments done under identical conditions. It should further be noted that a low silica recovery suggests a selectivity to the copper. The values given for copper recovery generally are correct to ±0.05 and those for silica recovery are generally correct to ±0.03. The data in Table I above clearly demonstrates the effectiveness of the collectors of the present invention. A comparison of Runs 1-5, not embodiments of the invention, with Runs 6-10 showed that at various pH levels, the monosulfonated collector of the present invention consistently resulted in substantially higher copper recoveries and comparable or lower silica recoveries. Example 2 - Flotation of Iron Oxide Ore A series of 600-g samples of iron oxide ore from Michigan were prepared. The ore contained a mixture of hematite, martite, goethite and magnetite mineral species. Each 600-g sample was ground along with 400 g of deionised water in a rod mill at about 60 RPM for 10 minutes. The resulting pulp was transferred to an Agitair 3000 ml flotation cell outfitted with an automated paddle removal system. The collector was added and the slurry allowed to condition for one minute. Next, an amount of a polyglycol ether frother equivalent to 40 g per ton of dry ore was added followed by another minute of conditioning. The float cell was agitated at 900 RPM and air introduced at a rate of 9.0 liters per minute. Samples of the froth concentrate were collected at 1.0 and 6.0 minutes after the start of the air flow. Samples of the froth concentrate and the tailings were dried, weighed and pulverised for analysis. They were then dissolved in acid, and the iron content determined by the use of a D.C. Plasma Spectrometer. Using the assay data, the fractional recoveries and grades were calculated using standard mass balance formulas. The results are shown in Table II following. A comparison of Runs 1 and 2 demonstrates that the use of the monosulfonated collector of this invention resulted in approximately a 50 percent increase in recovery of a slightly higher grade iron that is obtained using a disulfonated collector. Example 3 - Flotation of Rutile Ores A series of 30-g samples of a -10 mesh (U.S.) mixture of 10 percent rutile (TiO₂) and 90 percent silica (SiO₂) were prepared. Each sample of ore was ground with 15 g of deionised water in a rod mill - 2.5 inch (6.35 cm) diameter with 0.5 inch (1.27 cm) rods - for 240 revolutions. The resulting pulp was transferred to a 300 ml flotation cell. The pH of the slurry was left at the natural ore pH of 8.0. After addition of the collector as shown in Table III, the slurry was allowed to condition for one minute. Next, the frother, a polyglycol ether, was added in an amount equivalent to 0.050 kg per ton of dry ore and the slurry allowed to condition an additional minute. The float cell was agitated at 1800 RPM and air introduced at a rate of 2.7 liters per minute. Samples of the froth concentrate were collected by standard hand paddling at 1.0 and 6.0 minutes after the start of the introduction of air into the cell. Samples of the concentrate and the tailings were dried and analyzed as described in the previous examples. The results obtained are presented in Table III following. The data in Table III above demonstrates the effect of the collector of the present invention in increasing titanium grade and recovery. Comparison of Run 1 with Run 2 and Runs 4 and 5 with Run 3 again shows the marked improvements obtained using the monosulfonate collectors of this invention as compared to disulfonate collectors. Example 4 - Separation of Apatite and Silica A series of 30-g samples of a -10 mesh (U.S.) mixture of 10 percent apatite (Ca₅(Cl₁F)[PO₄]₃) and 90 percent silica (SiO₂) were prepared. The remainder of the procedure was exactly the same as that used in Example 3. The natural ore slurry pH was 7.1. In Runs 8-13, a blend of monosulfonated and disulfonated collector was used. The data in Table IV shows the ability of the process of this invention to separate apatite and silica. The information presented in Table IV demonstrates the marked effectiveness of the monosulfonated collectors in recovering phosphorus from an apatite and silica ore. Comparing Runs 2 and 4 to Runs 1 and 2, which were not examples of the invention, demonstrates the effect of monosulfonation. Runs 5-6 demonstrate that the collector of this invention was effective when used with an added hydrocarbon. A slight decrease in recovery was accompanied by a marked increase in grade. In Runs 8-13, the effect of mixing monosulfonated collectors and disulfonated collectors is demonstrated. A comparison of Runs 2, 11 and 13, wherein the levels of monosulfonated collectors are comparable and the amount of disulfonated species ranges from zero to 0.160 kg per metric ton, shows that the presence of the disulfonated species at low levels appeared to act as a diluent. At higher levels, the disulfonated species does not interfere with recovery, but does appear to lower the grade. Example 5 Samples (30 g of -10 mesh [U.S.]) of ore from Central Africa was prepared. The content of the copper metal in the ore was about 90 percent malachite with the remainder being other minerals of copper. Each sample of ore was ground along with 15 grams of deionized water in a mini-rod mill (2.5 inch diameter with 0.5 inch rods) for 1200 revolutions. The resulting pulp was transferred to a 300-ml mini-flotation cell. The pH of the slurry was left at a natural ore pH of 6.2. Collector was added at a dosage of 0.250 kg per metric ton of dry ore feed in Runs 1-20. In Runs 20-26, the collector dosage was varied and in Runs 22-26, the collector includes varying amounts of a disulfonate. After addition of the collector, the slurry was allowed to condition in the cell for one minute. Frother, a polyglycol ether, was added next at a dosage of 0.080 kg per metric ton of dry ore. This addition was followed by another minute of conditioning. The float cell was agitated at 1800 RPM and air introduced at a rate of 2.7 liters per minute. The froth concentrate was collected for 6.0 minutes. The samples of concentrates and tailing were then dried, weighed, pulverized for analysis and then dissolved with the use of acid. The copper content was determined by use of a D.C. plasma spectrometer. The information in the above table demonstrates the effectiveness of various alkylated diaryl oxide monosulfonates in the flotation of copper oxide ores. A comparison of the even numbered Runs 2-18 which are examples of the invention with the odd numbered Runs 1-19 which are not examples clearly demonstrates the substantially improved results obtained when using a monosulfonated collector as compared to a disulfonated collector when used at the same dosage. Comparing Run 2 with Run 21 demonstrates the effect of dosage. Runs 20-26 show that in blends, the disulfonated species appears to act as a diluent when blended with the monosulfonated collectors of this invention. Example 6 - Flotation of Iron Oxide Ore A series of 600-g samples of iron oxide ore from Michigan were prepared. The ore contained a mixture of hematite, martite, goethite and magnetite mineral species. Each 600-g sample was ground along with 400 g of deionized water in a rod mill at about 60 RPM for 15 minutes. The resulting pulp was transferred to an Agitair 3000 ml flotation cell outfitted with an automated paddle removal system. Flotation was conducted at the natural slurry pH of 7.0. Propylene glycol was added in the amount specified in Table VI below and the slurry allowed to condition for one minute. Next, the collector was added and the slurry allowed to condition for one minute. Next, an amount of a polyglycol ether frother equivalent to 40 g per ton of dry ore was added followed by another minute of conditioning. After comencement of flotation, additional collector was added in stages as shown in Table VI following. The float cell was agitated at 900 RPM and air introduced at a rate of 9.0 liters per minute. Samples of the froth concentrate were collected at intervals of zero to 1.0, 1.0 to 3.0, 3.0 to 4.0, 4.0 to 6.0, 6.0 to 7.0, 7.0 to 9.0, 9.0 to 10.0 and 10.0 to 14.0 minutes after the start of the air flow as shown in the table below. Samples of the froth concentrate and the tailings were dried, weighed and pulverized for analysis. They were then dissolved in acid, and the iron content determined by the use of a D.C. Plasma Spectrometer. Using the assay data, the fractional recoveries and grades were calculated using standard mass balance formulas. The results are shown in Table VI below. The data in Table VI above demonstrates that the monosulfonate collector of the present invention results in a very high recovery of high grade iron in substantially less time than comparable recoveries using the disulfonate. Example 7 - Flotation of Various Oxide Minerals The general procedure of Example 1 was followed with the exception that various oxide minerals were used in place of the copper ore. All runs were conducted at a pH of 8.0. The collector used was a branched C₁₂ dialkylated sodium diphenyl oxide monosulfonate at a dosage of 0.024 kg of collector per kilogram of mineral. The data in Table VII demonstrates the broad range of minerals which can be floated using the collector and process of this invention. Example 8 - Flotation of Mixed Copper Sulfide Ore Containing Molybdenum A series of 30-gram samples of a -10 mesh (U.S.) ore from Arizona containing a mixture of various copper oxide minerals and copper sulfide minerals plus minor amounts of molybdenum minerals were prepared. The grade of copper in the ore was 0.013 and the grade of the molybdenum was 0.000016. Each sample of ore was ground in a laboratory swing mill for 10 seconds and the resulting fines transferred to a 300 ml flotation cell. Each run was conducted at a natural ore slurry pH of 5.6. The collector was added at a dosage of 0.050 kg/ton of dry ore and the slurry was allowed to condition for one minute. Two concentrates were collected by standard hand paddling between zero and two minutes and two to six minutes. Just before flotation was initiated, a frother, a polyglycol ether available commercially from The Dow Chemical Company as Dowfroth® 250 brand frother, was added in an amount equivalent to 0.030 kg/ton of dry ore. The float cell in all runs was agitated at 1800 RPM and air introduced at a rate of 2.7 liters per minute. Samples of the concentrates and the tailings were then dried and analyzed as described in the previous examples. The results obtained are presented in Table VIII following. The data in Table VIII above demonstrates that the monosulfonated collector of the present invention obtains significantly improved recoveries of higher grade copper and molybdenum than does a comparable disulfonated collector. Example 9 - Hallimond Tube Flotation The procedure outlined in Example 1 was followed using a number of different mineral species and various collectors. Metal assays are performed on flotation concentrates and flotation tailings using acid dissolution and D.C. plasma spectrometry. The results are shown in Table IX following. While the data is presented in a single table, it is important to note that data on each mineral was obtained individually. In each instance the flotations were conducted at the natural pH of the respective ores in slurry form, i.e., 5.8 for rutile; 6.7 for apatite; 6.0 for pyrolusite; and 6.8 for diaspore. The data in Table IX above demonstrates that the monosulfonated collector used in the process of the present invention consistently obtains higher recoveries of a variety of minerals when compared to collectors that are similar other than for the monosulfonation. Example 10 - Sequential Flotation This example uses the Hallimond tube flotation procedure outlined in Example 1. In each case the feed material was a 50/50 percent by weight blend of the components listed in Table X below. The specific collectors used and the mineral recoveries obtained are also listed in Table X following. The data in the above table demonstrates that various minerals subject to flotation in the process of the present invention can be effectively separated by the control of collector dosage. For example, while apatite and dolomite can both be floated by the process of this invention, it is clear that apatite floats more readily at lower collector dosages than does dolomite. Thus, the apatite can be floated at a first stage, low dosage float. This can be followed by subsequent flotation at higher collector dosages to float the dolomite. As an examination of the other runs in this example demonstrate, similar separations are possible using other minerals. Example 11 - Separation of Apatite from Silica and Dolomite The procedure outlined in Example 4 was followed with the exception that the samples include 30 percent apatite, 60 percent silica and 10 percent dolomite. Additionally, a refined hydrocarbon was added in Runs 2 and 3. The results obtained are shown in Table XI following. The data in the above table demonstrates the ability of the collector of the present invention to float apatite preferably over dolomite or to separate apatite and dolomite. The industry standard shown in Run 4 does not obtain comparable separation of apatite and dolomite thus resulting in recovery of phosphorus significantly contaminated with magnesium. The addition of the hydrocarbon in the process of the present invention results in a slightly decreased recovery of higher grade phosphorus while decreasing the amount of magnesium collected. Example 12 - Flotation of Apatite The procedure followed in Example 11 was followed with the exception that the ore floated was a mixture of 30 percent apatite, 10 percent calcite and 60 percent silica. The results obtained are shown in Table XII following. The data in Table XII above demonstrates the effectiveness of the present invention in the recovery of apatite. When compared to Example 11, it also shows that the dosage needed to obtain a particular recovery is affected by the particular minerals being subjected to flotation. Example 13 - Flotation of Carbon Based Inks Five slurries were prepared by, in each case, pulping 240 g of printed paper (70 weight percent newsprint and 30 weight percent magazine); 1.61 g of diethylenetriaminepentaacetic acid, a color control agent; 10.65 g sodium silicate; the amount of the collector specified in Table XIII; and 5.64 g hydrogen peroxide with sufficient water to result in a slurry which was two weight percent solids. The slurry pH was 10.5, except as indicated and the temperature 45°C. Pulping was carried out for 30 minutes. Each slurry was prepared from copies of exactly the same pages to assure that the amount of ink was comparable in each of the five slurries prepared. The pulped slurry was transferred to a 15 liter Voith Flotation Cell with sufficient water of dilution to completely fill the cell. Sufficient calcium chloride was added to the pulp to give a water hardness of 180 parts per million CaCO₃. Flotation was initiated by the introduction of air bubbles passing through the highly agitated pulp and continued for a period of 10 minutes. Froth was then removed by standard hand paddling to produce the flotation product. The flotation product was then filtered and dried. The flotation cell contents containing the cellulose fibers were also filtered and dried. The flotation product was analyzed by colorimetry using a graded composition scale of 0 to 10 with 0 being all white and 10 being all black. The cellulose fiber mats prepared from the cell contents were examined using a high power microscope to observe the ink particles left per unit area. The data obtained is presented in Table XIII following. Conditions in each run are identical except as noted. The data in the above table demonstrates that the process of the present invention is effective in the separation of graphite ink and other carbon based inks from paper in the de-inking of recycled paper by flotation. Runs 2-5, when compared to Run 1 which approximates current industry standard, show that the use of the collectors of the present invention result in a greater recovery of ink at a significantly lower collector dosage.

Claims: A process for the recovery of minerals by froth flotation comprising subjecting an aqueous slurry comprising particulate minerals to froth flotation in the presence of a collector comprising an alkylated diaryl oxide sulfonic acid or salt thereof or a mixture of such acids or salts wherein at least 20 weight percent of the sulfonic acid or salts thereof are monosulfonated, under conditions such that the minerals to be recovered are floated. The process of Claim 1, wherein the monosulfonic acid or salt thereof corresponds to the formula: wherein each R is independently a saturated or unsaturated alkyl or substituted alkyl radical; each m and n is independently 0, 1 or 2 provided that at least one of m and n is not 0; each M is independently hydrogen, an alkali metal, alkaline earth metal, ammonium or substituted ammonium and each x and y are individually 0 or 1 with the proviso that the sum of x and y is 1. The process of Claim 2, wherein R is a linear or branched alkyl group. The process of Claim 3, wherein R is an alkyl group having from 1 to 24 carbon atoms. The process of Claim 4, wherein R is an alkyl group having from 6 to 24 carbon atoms. The process of Claim 5, wherein R is an alkyl group having from 10 to 16 carbon atoms. The process of any one of Claims 2 to 6, wherein the sum of m and n is two. The process of any one of the preceding claims, wherein the total concentration of the collector is at least 0.001 kg/metric ton and no greater than 5.0 kg/metric ton. The process of any one of the preceding claims, wherein at least 25 weight percent of the sulfonic acid or salt is monosulfonated. The process of any one of Claims 1 to 8, wherein at least 40 weight percent of the sulfonic acid or salt is monsulfonated. The process of any one of Claims 1 to 8, wherein at least 50 weight percent of the sulfonic acid or salt is monosulfonated. The process of any one of the preceding claims, wherein the recovered mineral comprises graphite and the aqueous slurry further comprises pulped paper. The use of an alkylated diaryl oxide sulfonic acid as defined in any one of Claims 1 to 11, as a froth flotation collector.

Inventors: KLIMPEL RICHARD ROBERT, LEONARD DONALD EUGENE, KLIMPEL, RICHARD ROBERT, LEONARD, DONALD EUGENE
Applicants: DOW CHEMICAL CO, THE DOW CHEMICAL COMPANY
Application year: 1990
Main classification: B03D1
IPC codes: C07C309, B03D1