PL163629B1 - Method of recovering minerals by froth flotation process - Google Patents

Abstract

1. Process for recovering minerals using froth flotation, which involves froth flotation of water suspension containing fragmented mineral with the presence of collector and possibly other flotation agents, generating conditions where the enriched minerals float to surface, characterised in that alkyl diaryl oxysulfonic acid or its salt or mixture of such acids or salts, with approximately 20% of sulfonic acid or its salts are monosulfonated.<IMAGE>

Description

The present invention relates to a method of recovering minerals by means of froth flotation. Flotation is the process of treating a mixture of finely ground mineral grains, e.g. powdered ore, suspended in a liquid, separating some of the solids from other, also ground, solid mineral materials, e.g. silica, siliceous gangue, clays and other similar materials found in an ore, by introducing gas (or generating gas in situ) into a liquid and creating a foamed mass containing some of the solid materials on the surface of the liquid and leaving other solid components of the ore in suspension (non-foamed). Flotation is based on the phenomenon that introducing gas into a liquid containing suspended solids of various materials causes some of the gas to stick to some of the suspended solids, which in turn causes the particles with gas attached to them to become lighter than the liquid. Therefore, these particles float to the surface of the liquid, creating foam.

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The minerals and associated gangue that are subject to froth flotation are generally not hydrophobic or hydrophilic enough to separate adequately. Therefore, various chemical reagents are used in the froth flotation process to impart or enhance the properties necessary for separation. Collectors are used to enhance the hydrophobic properties, and thus the flotation ability of various useful minerals. The collectors must adhere to the desired mineral species to the exclusion of other species present and maintain such connections under the turbulence and shear conditions associated with the froth flotation process, and must provide the desired mineral species with sufficient hydrophobic properties to achieve the desired degree of separation.

In addition to the collectors, several other chemical reagents are used, e.g. the type of 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 minerals and thus reduce their floatability. Frothers are added to the flotation systems to create a semi-permanent foam. Unlike depressants and collectors, frothers do not need to adhere to or absorb onto mineral particles.

Foam flotation has been used very often in the mining industry since at least the beginning of the 20th century. A wide variety of compounds are believed to be useful as collectors, frothers and other flotation reagents. In general, xanthogenates, simple alkylamines, alkyl sulfates, alkyl sulfonates, carboxylic acids, and fatty acids are included as useful collectors. Foaming reagents include low molecular weight alcohols such as methyl isobutyl methanol and glycol ethers. The specific additives to be used in a given flotation operation are selected depending on the nature of the rock, the conditions under which the flotation is carried out, the mineral to be recovered and other additives to be used with them.

A great number of chemicals are known to those skilled in the art that can be used in froth flotation, but it is also known that the effectiveness of known reagents varies considerably with the type of rock or rock to be flotated, as well as with the flotation conditions. It is also known that the selectivity or the ability to selectively flot the desired species with the exception of the undesirable species is a particular problem.

Minerals and their associated rocks are usually classified as sulfides or oxides, with oxides including oxygen-containing species such as carbonates, hydroxides, sulfates, and silicates. Thus, the group of minerals referred to as oxides usually includes oxygen-containing minerals. While a large proportion of naturally occurring minerals are contained in the oxide ores, the vast majority of satisfactory flotation systems are oriented towards sulphide ores. The flotation of oxide minerals is believed to be substantially more difficult than the flotation of sulfide minerals, and the effectiveness of most flotation processes in upgrading oxide ores is limited.

The main problem with upgrading both oxide and sulfide minerals is selectivity. Some of the known collectors, such as the alkyl carboxylic acids, sulfates, and sulfonates discussed above, are believed to be effective as collectors for oxide ores. However, while the use of these collectors may result in acceptable yields, it is believed that the selectivity for the desired minerals is quite poor, i.e. the degree or percentage of the desired component contained in the enriched mineral is unacceptably low.

As the degree of enrichment of oxide minerals obtained by conventional simple flotation is low, the mining industry tends to use more complex methods to achieve an acceptable degree of enrichment. In the industrial processes used, the oxide ores often undergo a sulfidation step prior to conventional flotation, followed by flotation using known sulfide collectors. Even with the sulfidation step, however, the yields and the degree of enrichment are less than desired. An alternative method of beneficiating oxide ores is liquid / liquid extraction. A third method used for the improvement of oxide ores, especially iron oxides and phosphates, is reverse or indirect flotation. In reverse flotation, the flotability of the ore containing the desired mineral decreases, as a result of which the waste rock and other impurities pass into

163,629 of the foam product. In some cases, the contaminant is a useful mineral. A fourth method of mineral enrichment involves chemical dissolution or leaching.

None of the currently known methods of flotation of oxide ores are without problems. In general, known methods result in a low yield and / or a low degree of enrichment. The low enrichment of the recovered minerals is considered to be a particular problem in the flotation of oxide minerals. Known beneficiation methods are not economical and consequently a large proportion of the oxide ores are simply not processed. There is therefore a great need for an oxide mineral flotation process with improved selectivity.

The method of mineral enrichment by means of froth flotation according to the invention consists in the froth flotation of an aqueous suspension containing particulate minerals in the presence of a collector containing diaryloxysulfonic acids or their salts or mixtures of such salts or acids, in which about 20% by weight are salts or monosulfonic acids, under conditions in which the recovered minerals flow to the surface. The recovered minerals may be useful minerals or undesirable contaminants. Additionally, frothers and other known flotation reagents are used in the froth flotation process of the invention.

In practice, the flotation process of the invention results in an improvement in selectivity and thus an improvement in the enrichment of minerals recovered from oxide and / or sulphide ores, typically while maintaining or increasing the overall yield of the desired mineral. It is surprising that the use of alkyldiphenyloximonosulfonated acids or their salts results in an improvement in the selectivity or the yield of useful minerals.

The flotation method of the present invention is suitable for recovering useful minerals from a variety of ores, including oxide ores, sulphide ores, and mixed ores.

Examples of the ores that can be flotated by the process of the invention preferably include oxides of iron, nickel, copper, phosphorus, aluminum and titanium. Other types of oxygen-containing minerals that can be flotated by the process of the invention include carbonates such as calcite or dolomite, and hydroxides such as bauxite.

Examples of specific oxide ores that can be flotated by the process of the invention include cassiterite, hematite, cuprite, valerite, calcite, talc, kaolin, apatite, dolomite, bauxite, spinel, corundum, laterite, azurite, rutile, magnetite, columbite, ilmenite, smitsonite, anglesite, shellite, chromite, cerisite, pyrolusite, malachite, chrysocolla, zincite, massicote, bixbyite, anatase, brucite, tungstite, uraninite, gummite, brucite, manganite, topical, topical, getite, limonite, chrysoberyl, chrysoberyl samarskite. The process of the invention is further applicable to the treatment of ores including oxide ores, wherein the term "oxide includes carbonates, hydroxides, sulfates, and silicates, and oxides.

The process of the invention is also suitable for the flotation of sulphide ores. Examples of sulfide ores which may be flotated by the process of the present invention include those containing chalcopyrite, chalcosine, galenite, pyrite, sphalerite, molybdenite and pentlandite.

Precious metals such as gold and silver, and metals from the platinum group of platinum, ruthenium, rhodium, palladium, osmium and iridium can also be recovered by the method of the invention. Such metals, for example, are sometimes associated with oxide and / or sulfide ores. Platinum, for example, can be found associated with troilite. By the process of the invention, such metals can be recovered in good yield.

The ores do not always exist as pure oxide or sulfide ores. Naturally occurring ores may contain both oxygen-containing minerals and sulfur-containing minerals, as well as small amounts of noble metals, such as those discussed above. The process of the present invention can also recover minerals from such mixed ores. This can be done by a two-step flotation, where one step involves conventional sulfide flotation to recover predominantly sulfide minerals, and the second step uses the method of the invention to recover mainly oxide minerals and precious metals that may be present. Alternatively, the sulfur-containing minerals and the oxygen-containing minerals can be recovered simultaneously by the process of the invention.

A special feature of the process according to the invention is the ability to flotate different minerals differently. It is believed that the flotation behavior of various minerals in the process of the invention is related to the crystalline structure of the minerals. Namely, it turns out that there is a correlation between the ratio of the crystal edge length to the crystal specific surface on the basis of the unit area. It turns out that minerals with higher ratios flotate more favorably than minerals with lower ratios. Thus, minerals with 24 or more faces (Group I) are usually more susceptible to flotation than minerals with 16 to 24 faces (Group II). Group III includes minerals that contain 12 to 16 faces and is next in preferential flotation susceptibility, followed by Group IV minerals with 8 to 12 faces.

In the process of the invention, generally Group I minerals will float before Group II minerals, which will in turn float before Group III minerals and those before Group IV minerals. By "pre-flotation" or preferential flotation is meant that the preferred grades will float at lower rates (amounts) of collector used. This means that Group I minerals can be harvested at a very low dose. After increasing the dose and / or removing most of the Group I minerals, Group II minerals will accumulate, etc.

It is understood by those skilled in the art that this classification of minerals is not the only one. Different minerals may have different possible crystal structures. In addition, the size of naturally occurring crystals also varies, which will affect the ease with which various minerals can flotate. An additional factor that affects the susceptibility to flotation is the degree of freedom. Further, within a group, i.e. among minerals whose crystals have similar edge-to-surface ratios, these and other factors will influence which group members will float first.

One skilled in the art can easily determine which group a mineral belongs to by standard examination of the mineralogical characteristics of various minerals. Such studies are, for example, reported in the Manual of Mineralogy, 19th ed., Cornelius S. Hurlbut, Jr. and Cornelis Klein (John Wiley and Sons, New York 1977).

Examples of Group I minerals include graphite, nicolite, covelite, molybdenite and beryllium.

Examples of Group II minerals include rutile, pyrolusite, anatas cassiterite, calomel, torbernite, autunite, marialite, meionite, apophyllite, zircon and xenothym.

Examples of Group III minerals include arsenic, grinokite, millerite, zincite, corundum, hematite, brucite, calcite, magnesite, siderite, rhodochrosite, smitsonite, soda niter, apatite, pyromophyte, mimetite, and vanadinite.

Examples of Group IV minerals include sulfur, chalcocyte, chalcopyrite, stybnite, bismuthinite, lelingite, marcasite, masycot, brukite, bemite, diaspor, getite, samarskite, atacamite, aragonite, witherite, strontium, cerisite, phosgenite, Indian nitrate, tenardite, barite , celestite, anglesite, epsomite anhydrite, anthlerite, caledomite, triphylite, lithiophyllite, heterosite, purpurite, varicite, strngite, chrysoberyl, shellite, desklizite, mottramite, brazilianite, olivenite, libetenite, adammite, children, ephouranylite, children, eshelite sulfite wulfenite, topaz, columbite and tantalite.

This classification into the groups discussed above is theoretical and carried out to identify which minerals will float preferentially. However, the method of the invention and the collector used therein are suitable for the flotation of various minerals that do not fit into the above categories. This cluster of minerals is useful for predicting which minerals will float at the relatively lowest collector rates, and not for determining what minerals at all may be harvested by the flotation method of the invention.

The selectivity of the collectors used in the process of the invention allows the separation of small amounts of undesirable minerals from useful minerals. For example, the presence of apatite is often a problem in iron flotation, as is the presence of topaz in cassiterite flotation. The collectors which are used in the process of the invention, in some cases, allow to perform a reverse flotation in which an undesirable mineral flows to the surface, such as cassiterite topaz or iron apatite.

In addition to the flotation of naturally occurring ores, the process of the invention is suitable for the flotation of minerals from other sources. One example is waste materials from various processes such as magnetic separation, metalworking and petroleum processing. These waste materials often contain minerals that can be recovered by the process of the invention.

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Another example is the recovery of a mixture of graphite ink and other carbon-based inks in the processing of newsprint. Typically such newspapers are discolored to separate inks from the paper fibers by flotation. The flotation method of the invention is particularly effective for such flotation paint removal.

The diaryloxymonosulfonic acid or monosulfonate which is a collector in the process of the invention is represented by the formula: Ar'-O-Ar where Ar 'and Ar are independently substituted or m-substituted aromatic groups, such as, for example, phenyl or naphthyl, provided that one and only one of the Ar 'and Ar groups contains one sulfonic acid or sulfonic acid salt group. Preferably, the diaryloxymonosulfonic acid or monosulfonate used as collector is an alkyldiphenyloxy- or alkylbiphenylphenyloximonosulfonic acid or a monosulfonate or a mixture thereof. The diaryloxymonosulfonic acid or monosulfonate is preferably substituted with one or more hydrocarbon substituents. The hydrocarbon substituents can be substituted or unsubstituted alkyl, or a substituted or unsubstituted unsaturated alkyl.

The monosulfonated diarylether collector in the process of the invention is particularly preferably a diphenylether collector which corresponds to the formula shown in the figure or a mixture of compounds represented by this formula, wherein each R independently is an optionally substituted saturated alkyl radical or an optionally substituted unsaturated alkyl radical, each with a different value. is independently 0, 1 or 2, each M is independently hydrogen, alkali metal, alkaline earth metal, ammonium or substituted ammonium, and each x and y is 0 or 1, with the proviso that the sum of x and y is 1. Preferably the group ( y) R is independently an alkyl group having 1 to 24, especially 6 to 24, carbon atoms, more preferably 6 to 16 and most preferably 10 to 16 carbon atoms. The alkyl groups may be linear, branched, or cyclic with linear or branched radicals being preferred. Preferably, each zin is 1. M + radicals representing an ammonium ion are represented by the formula (R ') 3NH +, wherein each R' independently is hydrogen, C1-C4 alkyl or C1-C4 hydroxyalkyl. These radicals include methyl, ethyl, propyl, isopropyl, butyl, hydroxymethyl and hydroxyethyl groups. Typical ammonium radicals include ammonium (N ⁺ H4), methyl ammonium (CHaN ⁺ H3), ethyl ammonium (C2 H ₅ N ⁺ H3), dimethyl ammonium (/ CH 3/2 N ⁺ H2), methyl ethyl ammonium (CH3N ⁺ H2C2Hs), trimethylammonium (/ CH 3/3 N ⁺ H), dimethylbutylammonium (/ CH 3/2 N ⁺ HC 4H9), hydroxyethylammonium (OHCHO2CH2N ⁺ H3) and methylhydroxyethylammonium (CH 3N ⁺ H 2CH 2CH 2O H).

Preferably, each M group is hydrogen, sodium, calcium, potassium or an ammonium group.

Alkylated diphenyloxysulfonates and methods for their preparation are well known in the art. The monosulfonate collectors used in the process of the invention can be made by modification of known methods for producing sulfonates. Representative methods for the preparation of sulfonates are known from U.S. Patent Nos. 3,264,242, 3,634,272, and 3,945,437. Commercial methods for preparing alkylated diphenyloxysulfonates generally do not produce compounds that are exclusively monoalkylated, monosulfonated, dialkylated or disulfonated. The commercially available compounds are mainly (greater than 90%) disulfonated and are a mixture of mono- and dialkylated compounds with 15 to 25% dialkylated and 75 to 85% monoalkylated. The most common commercial compounds are mixtures of about 80% monoalkylated and 20% dialkylated.

The use of monosulfonated compounds in the process of the invention has been found to be critical. Such monosulfonated compounds can be made by modifying the sulfonation step in a method described in, e.g., U.S. Pat. No. 3,264,242, 3,634,272 and 3,945,437. These methods are directed primarily at the preparation of disulfonated compounds. Hence, enough sulfonating agent is used in the sulfonation step to sulfonate both aromatic rings. In contrast, in the preparation of the monosulfonates which are useful in the process of the present invention, the amount of sulfonating agent used is limited to that necessary to provide one sulfonate group per molecule.

The monosulfonates so produced will contain both unsulfonated molecules and molecules that contain more than one sulfonate group per molecule. In case

The monsulfonates can be separated and used in relatively pure form. However, the mixture that results in a sulfonation step in which only an amount of sulfonating agent is used to provide approximately one sulfonate group per molecule is also suitable for use in the process of the invention.

As stated above, the use of monosulfonated compounds is critical to the process of the invention. However, the presence of disulfonated compounds is not considered to be deleterious in theory as long as at least 20% of the monosulfonated compounds are present. Preferably, the collector in the process of the invention comprises at least 25% monosulfonated compounds, more preferably at least 40% and most preferably at least 50%. The use of relatively pure monosulfonated acids or salts is most preferred. It will be appreciated by those skilled in the art that although high costs are associated with the production of relatively pure monosulfonated compounds in industrial use, this will be offset by the flotation efficiency compared to the reduced efficiencies associated with the use of mixtures containing disulfonated compounds.

Commercially available alkylated diphenyloxysulfonates are often mixtures of monoalkylated and dialkylated compounds. Although such mixtures are effective in the process of the invention, it is preferred, under certain conditions, to use compounds which are either monoalkylated, dialkylated or trialkylated. Such compounds are prepared by modification of methods described, for example, in U.S. Pat. No. 3,264,242, 3,634,272, and 3,945,437. If a system other than this mixture is desired, a distillation step is introduced after the alkylation step to remove the monoalkylated compounds and either monoalkylated compounds are used or recycled for further alkylation. In general, dialkylated compounds are preferred, although monoalkylated and trialkylated are also effective.

Examples of preferred alkylated diphenyl oxysulfonates include sodium monosulfonated diphenylether, sodium monosulfonated hexyldiphenyl ether, sodium monosulfonated ether for dodecyldiphenyl, sodium monosulfonated dodecyldiphenyl ether and sodium hexyldiphenyl monosulfonated ether of monosulfonated ether sodium salt. In a particularly preferred embodiment, the collector according to the invention uses the sodium salt of the monosulfonated dialkylated diphenyl ether in which the alkyl group is a group having 10-16 carbon atoms, most preferably 10-12 carbon atoms. The alkyl groups can be branched or linear.

In the process of the invention, the collector may be used at any concentration that gives the desired selectivity and yields the desired useful minerals. In particular, the concentration is dependent on the mineral being recovered, the richness of the ore to be flotated, and the desired quality of the mineral to be recovered.

Additional factors that should be considered in determining the dose level to be applied include the size of the specific surface area of the ore to be treated. It will be appreciated by one skilled in the art that with smaller particles, there is a larger specific ore surface area and a larger amount of collector is needed to obtain the desired yield and enrichment. Typically, the ores with the oxide minerals must be comminuted more than the sulfide ores and thus require a higher collector dose or removal of the finest particles by desludging. Conventional oxide mineral flotation processes typically require a desliming step to remove the finest particles and thus allow the process to run with an acceptable collector dose. In the method of the invention, the collector is operated at acceptable dose levels with or without desludging.

Preferably, the collector concentration is at least 0.001 kg / ton, more preferably at least 0.05 kg / ton. It is also preferred that the total collector concentration is not more than 5.0 kg / tonne, in particular not more than 2.5 kg / tonne. Usually for the most optimal performance it is best to start at low collector dosage levels and increase the level until the desired effect is achieved. In the method according to the invention, the yield and the degree of enrichment increase with increasing dose, and in some cases it will be understood by the skilled person that this increase is related to the increase in the cost of the flotation reagents. It is also understood that the dosages will also vary depending on the type of ore and other flotation conditions

163 629 collector. Moreover, it has been found that the collector dosages required are related to the amount of mineral to be collected. In those cases where small amounts of the mineral are amenable to flotation by the process of the invention, very small collector doses are needed due to the selectivity of the collector.

In enhancing certain minerals, it has been found advantageous to add the collector to the flotation system in steps. This should be understood to mean that a portion of the collector dose is added, the foam concentrate is collected, the next portion is added, and the foam concentrate is collected again. The total amount of collector preferably remains unchanged when added in stages. This stepwise addition can be repeated several times to obtain the optimal yield and degree of enrichment. The number of steps in which the collector is added is limited only by practical and economic conditions. Preferably no more than about 6 steps are used.

An additional advantage of staged addition relates to the collector's ability to flotate different minerals differently at different dose levels. As discussed above, at low dose levels, one mineral that is particularly flotable with the collector used in the present invention is flotated, while other minerals remain in the sludge. At an increased dosage, another mineral may be flotated and thus a separation of the various minerals contained in the ore is obtained.

In addition to the collector, other conventional reagents or additives may be used in the process of the invention. Examples of such additives include various known depressants and dispersants. In addition, it has been found advantageous to use hydroxyl containing compounds such as alkanolamines or alkylene glycols to improve selectivity to the desired minerals useful in silica or silicate gangue systems. In the process according to the invention, other collectors can be used with the above-defined collector. Frothers can also be used. Frothers are well known in the art. Examples of useful frothers include polyglycol ethers and lower molecular weight blowing alcohols.

A particular advantage of the collector used in the process of the invention is that no additional pH adjusting agents are needed for the flotation suspension. The flotation process of the invention using the collector as indicated above works efficiently at a typical natural rock pH in the range from 5 or below to 9. This is especially important when considering the cost of the reagents needed to adjust the pH of the slurry from a natural pH of about 7.0 or below to 9.0 or 10.0 or above, which is required in conventional methods using conventional carboxyl, sulfone, phosphono and xanthene collectors.

The ability of the collector used in the process of the invention to operate at a relatively low pH means that it can also be used in those cases where it is desired to lower the pH of the suspension. The lower pH limit of the slurry at which the process of the invention can be carried out determines the pH at which the surface charge on the mineral particles is suitable for attachment to the collector.

Since the collector in the method of the invention operates at different pH values, the tendency of different minerals to float at different pH levels can therefore be used. This enables one flotation operation to be carried out at one pH value in order to optimize the flotation of a given mineral. The pH can then be adjusted to the next flotation to optimize the flotation of the other component, which facilitates the separation of the different minerals present together.

In the method of the invention, the collector can be used together with other conventional collectors. For example, monosulfonated diaryl ether collectors can be used in a two-stage flotation in which the ether collector is mainly enriched with oxide minerals in a first step, and in a second step, using conventional collectors, mainly sulfide minerals or other oxide minerals are enriched. When conventional collectors are used in the two-stage flotation, the first stage can be carried out by the process of the invention at the natural pH of the slurry. The second step is carried out at elevated pH using conventional collectors. Note that in some cases you may need to reverse the order of the steps. Such a two-step process has the advantage of being applicable

With fewer additives to regulate the pH, it allows a more complete enrichment of the desired minerals by carrying out flotation under various conditions.

The following examples illustrate the invention but do not limit its scope. All parts and percentages are by weight unless otherwise stated.

The following examples include Hallimonda tube flotation and lab-scale flotation flotation. It should be noted that the flotation in the Hallimonda pipe is a simple method of selecting collectors, but on its basis it cannot be concluded that the collectors work positively in the actual flotation. Hallimonda pipe flotation does not experience the shear or mixing that occurs in true flotation, and the effect of foaming agents is not measured. A collector therefore needs to be effective in Hallimonda tube flotation if it is to be effective in true flotation, while a collector that is effective in Hallimonda flotation need not be effective in true flotation. It should also be noted that experiments show that the collector doses required to obtain a satisfactory yield in the Hallimonda pipe are often much higher than the doses required in the trials in flotation machines. Therefore, based on the work carried out in the Hallimonda pipe, it is not possible to accurately predict the doses to be used in a real flotation machine.

Example I. Flotation of malachite and silica in a Hallimonda pipe.

About 1.1 g of (1) malachite, an oxide copper mineral with the approximate formula Cu2CO3 (OH) 2, or (2) silica is ground into grains ranging from about -60 to +120 mesh USA and placed in a small bottle along with with 20 ml of deionized water. The mixture was shaken for 30 seconds, after which the aqueous phase, containing some suspended fine particles and sludge, was decanted. This desliming step was repeated several times.

150 ml of deionized water was placed in a 250 ml glass beaker. Then 2.0 mL of 0.10 M potassium nitrate solution was added as buffer electrolyte. The pH was adjusted to about 10.0 by adding 0.10 N HCl and / or 0.10 N NaOH. A 1.0 g portion of the desludged mineral was then added along with deionized water to bring the total volume to about 180 ml. A collector was added which was a C16 alkylated sodium diphenyloxysulfonate containing about 80% monoalkylated compound and about 20% dialkylated compound, followed by conditioning with stirring for 15 minutes. The pH was monitored and adjusted with HCl and NaOH as necessary. It should be noted that Runs 1-5, which use a disulfonated collector, do not represent the invention, while Runs 6-10, which illustrate the method of the invention, use a monosulfonated collector. This is the only difference in the collectors used in trials 1-5 and 6-10.

The slurry was transferred to a Hallimond tube made so that a capillary could be placed at the bottom of the 180 ml tube. After adding the slurry to the Hallimond tube, a vacuum of 17 kPa was applied to the open end of the tube for 10 minutes. This negative pressure caused air bubbles to enter the tube through the capillary inserted into the bottom. During flotation, the slurry was stirred with a magnetic stirrer set at 200 rpm. ·

The floated and non-floated material was filtered off the slurry and dried in an oven at 100 ° C. Each batch was weighed and the partial yields of copper and silica are shown in Table 1. After each run, the entire unit was washed with concentrated HCl and rinsed with 0.10 N NaOH and deionized water before the next run.

The copper and silica yield is represented as a fraction of the starting mineral placed in the Hallimonda pipe. Therefore, a yield of 1.00 means that all material has been recovered. Note that although the yields of copper and silica are given together, the data were obtained from two experiments run under identical conditions. It should also be noted that the low yield of silica suggests selectivity over copper. The values given as the copper yield are given with an accuracy of ± 0.05, and the silica yield with an accuracy of ± 0.03.

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Table 1

Fractional share of the yield

Attempt Collector Dose pH Cu silica 1 (2) LC, e -DPO (SO ₃ Na) 2 (1) 0 060 5 5 0 760 0 153 2 (2) LC, 6 -DPO (SO ₃ Na) 2 (1) 0 060 70 0 809 0.082 3 (2) LC, _e -DPO (SO ₃ Na) 2 (1) 0 060 8 5 0 800 0 062 4 (2) LC, e-DPO (SO ₃ Na) 2 (1) 0 060 10 0 0 546 0 104 5 (2) LC, 6-DPO (SO ₃ Na) 2 (1) 0 060 11 5 0 541 0 130 6 LC, e -DPO (SO ₃ Na) 1 (3) 0 060 5 5 0 954 0 135 7 LC, 8-DPO (SO ₃ Na) 1 (3) 0 060 7.0 0 968 0 097 8 L-C, 6-DPO (SO3Na) 1 (3) 0 060 85 0 913 0 840 9 L-Cie -DPO (SO3Na), (3) 0 060 100 0 837 0 070 10 L-C, a -DPO (SO3Na) 1 (3) 0 060 11 5 0 798 0 065

(1) - Linear Cie alkylated diphenyloxysulfomate sodium salt, contains about 80% mono- and 20% dialkylated compounds, commercially available from The Dow Chemical Company as a surfactant called DOWFAX 8390 (2) - not according to the invention (3) - Linear Cie alkylated diphenyloximonosulfonate sodium salt, 80% monoalkyl and 20% dialkylated

The results presented in Table 1 clearly show the efficiency of the collector used in the process of the invention. A comparison of trials 1-5, which do not illustrate the process of the invention, with trials 6-10, shows that at different pH levels the monosulfonated collector used in the process of the invention consistently gives substantially higher copper yields and comparable or lower silica yields.

Example II. Iron oxide ore flotation.

A series of 600 g samples of Michigan oxide iron ore were prepared. The ore was a mixture of hematite, martite, goetite and magnetite. Each of the 600g samples was milled with 400g deionized water for 10 minutes in a rod mill rotating at approximately 60 rpm. The obtained pulp was transferred to a 3000 ml Agitair flotation machine equipped with an automated paddle-type foam removal system. A collector was added and the slurry conditioned for 1 minute. Subsequently, polyglycol ether was added as a foaming agent in an amount corresponding to 40 g / tonne of dry ore, followed by conditioning for 1 minute.

The flotation cell was mixed at a speed of 900 revolutions / minute, introducing air into it in the amount of 9.01 / minute. Samples of the foam concentrate were collected 1 and 6 minutes after the start of the air flow. The foam concentrate and waste samples were dried, weighed, and ground for analysis. They were then dissolved in acid and the iron content was determined using a D.C. plasma spectrometer. The results of the analysis were used to determine the fractional yields and the enrichment ratios using the usual mass balance equations. The results are presented in Table 2.

Table 2

The yield and degree of iron enrichment

Attempt Collector Dose 0-1 minutes 1-6 minutes Together yield st wzb yield st wzb yield st 1 (1) R D-C, 2DPO (SO3Na) 2 (2) 0 200 0 494 0 462 0 106 0 394 0 600 0 450 2 R D-C, 2DPO (SO3Na) 1 (3) 0 200 0 677 0 487 0 240 0 401 0.907 0 464

(1) - not according to the invention (2) - C12 branched alkyl diphenyloxymonosulfomate sodium salt (3) - C12 branched alkyl diphenyloximonsulfonate sodium salt

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Comparison of trials 1 and 2 shows that the use of a monosulfonated collector gives approximately a 50% increase in iron yield with a slightly higher enrichment than the results obtained with the disulfonated collector.

Example III. Rutile ores flotation.

A series of 30g samples were prepared with a -10 mesh (USA) mixture of 10% rutile (TiO ₂ ) and 90% silica (SiO 2). Each ore sample was milled with 15 grams of deionized water in a bar mill with 1.27 cm long, 6.35 cm diameter rods rotated at 240 revolutions. The obtained pulp was transferred to a 300 ml flotation machine. The pH of the slurry was left at the natural pH of the ore, i.e. 8.0. After adding the collector shown in Table 3, the slurry was conditioned for 1 minute. The frother polyglycol ether was then added in an amount corresponding to 0.050 kg / tonne of dry ore and the slurry was conditioned for a further minute.

The flotation cell was agitated at a speed of 1800 rpm, introducing air into it in the amount of 2.7 l / minute. Samples of the foam concentrate were collected 1 and 6 minutes after the start of the air flow. Samples of the foam concentrate and waste were dried and analyzed as described in the previous examples. The obtained results are presented in Table 3.

Table 3

A mixture of rutile and silica

Attempt Collector Dose Titanium yield and enrichment 0-1 minutes 1-6 minutes Together yield st wzb yield st. yield st. 1 (i) L D-C, 6-DPO (SOjNa) 2 (2) 0 200 0 677 0 086 0 061 0 064 0.738 0 084 2 L DC, 6-DPO (SO ₃ Na) 1 (3) 0 100 0 763 0 110 0 151 0 074 0.914 0 104 3 (1 R D-C12-DPO (SO ₃ Na) 2 (4) 0 200 0 756 0 099 0 086 0 075 0.842 0 097 4 R D-C12-DPO (SO ₃ Na) 1 (5) 0 200 0 809 0 077 0.134 0 066 0 943 0 075 5 R D-Cu-DPO (SO ₃ Na) 1 (5) 0 100 0714 0 086 0 117 0 070 0 831 0 084 6 R D-Cu) -DPO (SO ₃ Na) 1 (6) 0 100 0 674 0 095 0 099 0 071 0 773 0.092

Table 3 (references).

(1) not according to the invention (2) sodium salt of diphenyloxydisulfonate dialkylated with C16 linear alkyl (3) sodium salt of dialkylated diphenyloximonsulfonate dialkylated with linear alkyl Ct6 (4) sodium salt of dialkylated branched alkyl diphenyloxydisulfonate (5) branched alkyl sodium salt C12 dialkyloximonate) C10 branched alkyl diphenyloximonosulfonate sodium.

The results shown in Table 3 above show the increase in the yield and the degree of titanium enrichment in the process of the invention. Comparing Run 1 to Run 2, and Runs 4 and 5 to Run 3 show the significant advantages obtained by using the monosulfonate collectors in the process of the invention over the disulfonate collectors.

Example IV. Separation of apatite and silica.

A series of 30g samples with -10 mesh grindings of a mixture of 10% apatite (Ca5) Cl1F ([PC> 4] 3) and 90% silica was prepared. The further procedure was exactly the same as in Example III. The natural ore slurry has a pH of 7.1. Runs 8-13 used a mixture of monosulfonated and disulfonated collectors. The results presented in Table 4 demonstrate the possibility of separating apatite and silica by the process of the invention.

Table 4

Attempt Collector Dose. kg / Mg P. Yield Degree enrichment 1 (1) L, D-C, 0DPO (SO3Na) 2 (2) 0 050 0 115 0.081 2 L.D-C, 0DPO (SO3Na) 1 (3) 0 050 0 962 0.068

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1 2 3 4 5 3 (1) R, D-C1 ₂ DPO (SO3Na) ₂ (4) 0 050 0 235 0 078 4 R, DC ₂ DPO (SO3Na) 1 (5) 0 050 0 989 0 067 5 L, D-C1oDPO (SO3Na), (3) 0 050 0 925 0 103 Refined kerosene (6) 0 500 6 Refined kerosene (6) 0010 L, D-C, oDPO (SO3Na) 1 (3) 0 050 0.862 0 112 7 Refined kerosene (6) 0 020 L.D-Ck, DPO (SO3 Na) 1 (3) 0 050 0 818 0 125 8 L.D-C10DPO (SO3Na) 2 (2) 0 040 L.D-C, oDPO (SO3Na) 1 (3) 0010 0 336 0 077 9 L, D-C, 0DPO (SO3Na) 2 (2) 0 030 L.D-C, 0DPO (SO3Na) 1 (3) 0 020 0 529 0 075 10 L.D-C, 0DPO (SO3Na) 2 (2) 0 020 L, D-C, 0DPO (SO3Na) 1 (3) 0 030 0 699 0 074 11 L.D-C, 0DPO (SO3Na) 2 (2) 0010 L, D-C10DPO (SO3Na), (3) 0 040 0 866 0 069 12 L.D-C, 0DPO (SO3Na) 2 (2) 0 080 L.D-C, 0DPO (SO3Na) 1 (3) 0 020 0 539 0.067 13 L.D-C, 0DPO (SO3Na) 2 (2) 0 160 L, D-C, oDPO (S03Na), (3) 0 040 0 877 0 53

Table 4 (references) (1) Me according to the invention (2) C10 linear alkyl dialkylated diphenyloxydisulfonate (3) Linear alkyl dialkylated sodium diphenyloximonsulfonate

C10 (4) C12 branched alkyl dialkyloxymonosulfonate sodium salt (5) C12 branched alkyl dialkylate sodium salt (6) Phillips Petroleum refined petroleum product Soltrol, added simultaneously with the collector to the flotation chamber

The results presented in Table 4 demonstrate the clear efficiency of monosulfonated collectors in recovering phosphorus from apatite and silica ore. A comparison of trials 2 and 4 with trials 1 and 2, which do not illustrate the process of the invention, shows the performance of the monosulfonates. Trials 5-6 illustrate the fact that the manifold used in the process of the invention was effective when used together with the added hydrocarbon. There was a slight decrease in the yield with a significant increase in the degree of enrichment. Trials 8-13 show the mixing effect of monosulfonated and disulfonated collectors. A comparison of trials 2, 11 and 13, in which the levels of monosulfonated collectors are comparable and the amounts of disulfonated compounds are in the range of 0 to 0.160 kg / ton, shows that the low levels of disulfonated compounds act as a diluent. In higher amounts, the disulfonated compounds do not affect the yield but reduce the enrichment degree.

ExampleV. Ore samples from Central Africa (30g, 10 mesh) were prepared. The metallic copper content of the ore was about 90% in malachite and the rest in other copper minerals. Each ore sample was milled with 15 g of deionized water in a mini-bar mill (1.77 cm long, 6.35 cm diameter rods) at a speed of 1200 revolutions. The produced slurry was transferred to a 300 ml miniflower. The pH of the slurry remained at the pH of the natural ore of 6.2. For runs 1-20, the collector was added at a rate of 0.250 kg / tonne dry ore. In Runs 20-26, the collector dose was varied and in Runs 22-26, the collector contained varying amounts of disulfonate. After the collector was added, the slurry was conditioned in the chamber for 1 minute. Then the frother - polyglycol ether was added at a dose of 0.080 kg / tonne of dry ore. Then it was conditioned for a further minute.

The machine was mixed at a speed of 1800 rpm and air was introduced at a speed of 2.7 l / min. The foam concentrate was collected in 6 minutes. Then the concentrate and waste samples were dried, weighed, pulverized for analytical purposes and dissolved with acid. The copper content was determined using a D.C. plasma spectrometer.

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Table 5

Attempt Collector Dose kg / Mg pH Cu Yield degree enrichment 1 2 3 4 5 6 1 (1) - - 62 0 038 0019 2 R, D-C1 ₂ DPO (SO ₃ Na), (2) 0 250 62 0 696 0 057 3 (1) R, D-C12DPO (SO ₃ Na) 2 (3) 0 250 62 0 501 0 042 4 LD-C, 0DPO (SO ₃ Na) 2 (4) 0 250 62 0 674 0 056 5 (1) L, D-Cu) DPO (SO ₃ Na) 2 (5) 0 250 62 0 487 0 035 6 L, D-CioDPO (SO3Na) 1 (6) 0 250 62 0 696 0 59 7 (1) L, D-C, 0DPO (SO3Na) 2 (7) 0 250 62 0 573 0 051 8 L ^ -tC ^ PO ^ Na), (8) 0 250 62 0 714 0 058 9 (1) L, D-C, 6DPO (SO3Na) 2 (9) 0 250 62 0 598 0 052 10 L.M-C, 0DPO (SO3Na) 1 (10) 0 250 6 2 0 390 0 046 11 (1) L, MC _{1 0} DPO (SO3Na) 2 (11) 0 250 62 0 116 0 038 12 R.M-C12DPO (SO3Na) 1 (12) 0 250 62 0 338 044 13 (1) R, M-C1 ₂ DPO (SO3Na) 2 (13) 0 250 62 0 145 0 041 14 L, lM-C24DPO (SO3Nal1 (14) 0 250 62 0 474 0 037 15 (1) L, M-C24DPO (SO3Na) 2 (15) 0 250 62 0 335 0 035 16 L M-CeDPO (SO3Nal1 (16) 0 250 62 0 111 0 037 17 (1) L, MC _e DPO (SO3Na) 2 (17) 0 250 62 0 053 0.038 18 L, D-CeDPO (SO3Na1l (18) 0 250 62 0317 0 041 19 (1) L, D-CeDPO (SO3Nal2 (^ 91 0 250 6.2 0 198 0.38 twenty R, D-C12DPO (SO3Na) 1 (2) 0 400 6 2 0.839 0.055 21 (1) R, D-C12DPO (SO3Na) 2 (3) 0 400 62 0 533 0 039 22 R, D-C12DPO (SO3Na) 1 (2) 0 I00 62 0 620 0 045 R, D-C12DPO (SO3Na) 2 (3) 0 300 23 R, DC, ₂ DPO (SO3Na) 1 (2) 0 200 62 0 683 0 051 R, DC, ₂ DPO (SO3Na) 2 (3) 0 200 24 R, D-C12DPO (SO3Na), (2) 0 300 62 0 788 0 054 R, D-C, 2DPO (SO3Na) 2 (3) 0 100 25 R, D-C12DPO (SO3Na) t (2) 0 400 62 0 855 0 041 R.D-C12DPO (SO3Na) 2 (3) 0 400 26 R, D-C12DPO (SO3Na) and (2) 0 400 62 0 861 0.039 R.D-C12DPO (SO3Na) 2 (3) 1,200

(1) not according to the invention (2) C10 branched alkyl diphenyloxymonosulfomate sodium salt (3) C12 branched alkyl diphenyloxydisulfonate sodium salt (4) C10 linear alkyl diphenyloximonsulfomate dialkylated sodium salt (5) C10 linear alkyl dialkyl sulfonate sodium di-dialkyl sulfonate 6 ) C10 linear alkyl biphenylphenyl oximonosulfomate dialkylated C10 (7) C10 linear alkyl biphenylphenyloxydisulfonate (8) linear alkyl diphenyloximonosulfomate dialkylated sodium salt C16 (9) sodium diphenyloxydisulfonate sodium dialkyl sulfonate sodium salt C18 linear alkyl oxime sulfonate sodium salt C18 linear alkyl oxime sulfonate C10 linear alkyl monoalkylated (12) C12 branched alkyl monoalkylated sodium salt (13) C12 branched alkyl monoalkylated sodium diphenyloxydisulfonate (14) C12 branched alkyl diphenyloxydisulfomate sodium salt m monoalkylated with linear alkyl C24 (15) sodium salt of monoalkylated diphenyloxydisulfonate monoalkylated with C24 linear alkyl (16) sodium salt of diphenyloximonosulfomate monoalkylated with linear alkyl C6 (17) sodium salt of monoalkylated diphenyloxydisulfonate monoalkylated with linear alkyl C8 (18) sodium dialkyl sulfonate diphenylsulfonate sodium diphenyl oximonate 19 dialkylated with linear Ce alkyl

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The results in the table above demonstrate the effectiveness of the various alkylated diaryloxymonosulfonates in the flotation of copper oxide ores. The comparison of the even-numbered trials 2-18, which are examples of the inventive process, with the odd-numbered trials 1-19, which are not examples of the inventive process, clearly show the substantially better results obtained with the monosulfonated collector compared to the disulfonated collector used in this invention. same dose. Comparison of trials 2 and 21 illustrate the dose effect. From trials 2-26 it appears that when mixtures of collectors are used the disulfonated compounds act as a diluent.

Example VI. Iron oxide ore flotation

A series of 600 g samples of Michigan iron oxide ore were prepared. The ore contained a mixture of hematite, goetite and magnetite. Each of the 600 g samples was milled with 400 g deionized water in a rod mill at approximately 60 rpm for 15 minutes. The slurry produced was transferred to a 3000 ml Agitair flotation machine equipped with an automatic scooping system. Flotation was performed at the natural pH of the suspension of 7.0. Propylene glycol was added in the amount specified in Table 6 and the suspension was conditioned for 1 minute. The collector was then added and the slurry was further conditioned for 1 minute. The polyglycol ether frother was then added in an amount equivalent to 40 g / tonne of dry ore, followed by conditioning for a further minute. After the flotation was started, the collector was additionally added in steps as shown in Table 6.

The flotage cell was agitated at 900 rpm and air was introduced at a rate of 9.0 L / min. Samples of the foam concentrate were collected from 0 to 1 minute, 1 to 3 minutes, 3 to 4 minutes, 4 to 6 minutes, 6 to 7 minutes, 7 to 9 minutes, 9 to 10 minutes, and 10 to 14 minutes after the start of the flow air as shown in the table below. Samples of the foam concentrate and waste were dried, weighed and pulverized for analytical purposes. They were then dissolved in acid and the iron content was determined with a D plasma spectrometer. Using the test results, the fractional yields and enrichment ratios were calculated based on the standard mass balance equation. The results are presented in Table 6.

Table 6

Collector Dose Iron yield and degree of enrichment Together Degree Enrichment yield Degree Enrichment yield Yield Degree enrichment 0-1 minute 1-3 minutes Glycol 0.100 0.112 0.514 0.028 0.461 0.140 0.503 B, D-C12-DPO (SO ₃ Na) ₂ (1) (2) 0.042 3-4 minutes 4-6 minutes B, D-C1 ₂ DPO (SO ₃ Na) 2 (1) (2) 0.042 0.231 0.538 0.061 0.550 0.432 0.528 6-7 minutes 7-9 minutes B, DC, 2D PO (SO ₃ Na) ₃ (1) (2) 0.042 0.178 0.488 0.045 0.493 0.655 0.515 9-10 minutes 10-14 minutes B, D-C1 ₃ DPO (SO ₃ Na) 2 (1) (2) 0.042 0.094 0.366 0.096 0.283 0.845 0.472 Propylene glycol 0-1 minute 1-3 minutes B, DC, 2DPO (SO ₃ N3) 2 (3) 0.100 0.353 0.526 0.042 0.157 0.498 0.510 0.517 3-4 minutes 4-6 minutes BD-C, 2DPO (SO ₃ Na) 1 (3) 0.042 0.219 0.508 0.099 0.487 0.828 0.510

(1) - not according to the invention (2) - C12 branched alkyl dialkylated sodium salt (3) C12 branched alkyl dialkylated sodium salt

The data in Table 6 show that the monosulfonate collector used in the process of the invention gives a much higher yield with a high degree of iron enrichment in substantially less time than comparable yields obtained with the disulfonates.

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Example VII. Flotation of various oxide minerals.

The general procedure was that of Example 1, except that various oxide minerals were used in place of the copper ore. All tests were performed at pH 8.0. The collector used was C12 branched alkyl sodium diphenyloximonosulfonate at a rate of 0.024 kg collector / kg of mineral.

This bale 7

Mineral Fractional yield Silica (SiO2) 0 204 Cassiterite (SnO2) 0 931 Bauxite [Al (OH 3)] 0 989 Calcite (CaCOa) 0 957 Chromite (FeCr2O4) l,000 Dolomite [CaMg (COs) 2] 0 968 Malachite [Cu2COs (OH2)] 0 989 Chrysocolla [Cu 2H 2Si2Os (OH) ₄ ] 0 616 Hematite (Fe2O3) 0 971 Corundum (AUO3) 1,000 Rutile (TiO 2) 0 970 Apatite [Ca 5 (Clif) [PO _4] 3 0 990 Nickel Oxide (NiO) 0 778 Galena (PbS) 0 990 Chalcopyrite (CuFeS 2) 0 991 Chalcocyte (CU2S) 0 993 Pyrite (FeS2) 1,000 Sphalerite (ZnS) 1,000

(1) The sample contains some pyrotite (2) The sample is elemental metal powdered with minerals

Minera! Fractional yield Pentlandite [Ni (FeS)] (1) 0 980 Metallic Cu (2) 0 931 Metallic Au (2) 0 964 Metallic Ag (2) 0 873 Baryte (BaSO ₄ ) 0 968 Molybdenum (M 0S2) 0 968 Cerusite (PbCO3) 0 939 Calcite (CaC03) 0 807 Beryl (Be3Al2SieO1e) 0 937 Kovelin (CuS) 0 788 Zirconium (ZrSiO4) 0 876 Graphite (C) 0 937 Topaz [AlaSiO 4 (F, OH) 2] 0.955 Shelit (CaWO4) 0.871 Anatase (TiO 2) 0 909 Bemit (yAl OH) 0 886 Diaspor (aAlOOH) 0 905 Gethyte (HFeO2) 0 959

a particle size similar to that of other samples

The results in Table 7 show the wide range of minerals that can be flotated by the process of the invention.

Example VIII. Flotation of a mixed copper sulphide ore containing molybdenum.

A series of 30g samples were prepared with a fineness of 10 mesh Arizona ore containing a mixture of various copper oxide minerals and copper sulphide minerals and smaller amounts of molybdenum minerals. The copper content of the ore is 0.013 and the molybdenum 0.000016.

Each ore sample was milled in a rotary mill for 10 seconds and the resulting finely divided sample was transferred to a 300 ml flotation machine.

Each test was carried out at the natural pH of the ore slurry of 5.6. The collector was added at a rate of 0.050 kg / tonne dry ore and the slurry conditioned for 1 minute. Between 0 and 2 minutes and between 2 and 6 minutes, the two concentrates were collected with standard hand scoops. Just before the start of the flotation, a frother, polyether glycol under the trade name Dowfroth 250 from The Dow Chemical Company, was added in an amount corresponding to 0.030 kg / tonne dry ore.

The flotation machine in all tests was mixed at a speed of 1800 rpm and air was introduced at a speed of 2.7 l / min. The concentrate and waste samples were dried and analyzed as described in the previous examples. The obtained results are presented in Table 8.

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Table 8

0-2 minutes

2-6 minutes

Collector Dose. kg / ton Cu Cu Mo Mo Cu Cu Mo Mo Uzvsk Degree enrichment Yield Degree enrichment Yield Degree enrichment Yield Degree enrichment L D-C, o DPOfSOjNa), (1) 0.050 0 820 0 169 0.875 000042 0.850 0.088 0 042 000011 L.D-C, o DPCHSOjNate (2) 0 050 0 447 0.133 0 706 000025 0.151 0 116 0.039 000005 L.D-Cio DPOfSOjNa). (3) 0.025 0 533 0.148 0.77 | 000026 0.232 0.130 0 041 000003

Table 8 (continued) Collector Dose kg / ton Total yield and degree of upgrading Cu Yield Cu Degree enrichment Mo Yield Mo Degree enrichment L.D-CwDPOtSOjNa). (1) 0 050 0.905 0 161 0.917 000040 L.D-C, <0DPO (SOjNab (2) 0.050 0.598 0.129 0.745 000024 L.D-CwDPCHSOjNab (1) 0.025 0.765 0.143 0.812 000025

(1) - sodium salt of diphenyloxymonosulfonate dialkylated with branched alkyl Cu (2) - sodium salt of diphenyloxydisulfoman dialkylated with branched alkyl C ia

The results presented in Table 8 show that the monosulfonated collector produces much better yields with a higher degree of copper and molybdenum enrichment than the comparable disulfonated collector.

Example IX. Hallimonda tube flotation.

The procedure described in Example 1 was followed using several different species of minerals and different collectors. An analysis of the concentrates and flotation waste was performed using acid dissolution and D.C plasma spectrometry. The results are shown in Table 9. The results for each mineral were obtained separately. In each case, the flotation was carried out at the natural pH of the corresponding slurry ore, i.e. at pH 5.8 for rutile, 6.7 for apatite, 6.0 for pyrolusite and 6.8 for diaspor.

Table 9

Attempt

Collector

Yield

kg / kg Rutile Apatite Pyrolusite Diaspor · 1 2 3 4 5 6 7 1 B, D-C1a DPO (SOaNa) 1 (1) 0 0001 0 021 0 009 - - 2 B, D-C, and DPO (SOaNa), (1) 0 0005 0 323 0 038 - - 3 BD-Ca DPO (SO ₃ Na) 1 (1) 0 0010 0713 0 463 - - 4 B.D-C1 DPO (SOaNa) 1 (1) 0 0100 0 954 0 856 0 745 598 5 (2) B, DC.a DPO (SO ₃ Na) with (3) 0 0001 0,000 0,000 - - 6 (2) B, D-C1a DPO (SO ₃ Na) a (3) 0 0005 0015 0 007 - - 7 (2) B, D-C1a DPO (SOaNa) ₂ (3) 0 0010 0 087 0 297 - - 8 (2) B, D-C1a DPO (SO ₃ Na) a (3) 0 0100 0 175 0 518 0314 0.280 9 (2) B, D-C1a DPO (SO ₃ Na) 2 (3) 0 0500 0 371 - - - 10 (2) B, D-C1a DPO (SO ₃ Na) ₂ (3) 0 1000 0815 0.849 - - II B, M-Cn. DPO (SO ₃ Na) 1 (4) 0 0001 0,000 0,000 - - 12 B, M-C1a DPO (SO ₃ Na) 1 (4) 0 0005 0011 0,000 - - 13 B, MC, a DPO (SO ₃ Na) 1 (4) 0.0010 0.034 0.111 - - 14 B, M-Cu DPO (SO ₃ Na) 1 (4) 0 0100 0 129 0 277 0 289 0.166

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1 2 3 4 5 6 7 15 B, MC, 2 DPO (SO ₃ Na), (4) 0 0500 0 296 - - - 16 B, MC, 2 DPO (SO ₃ Na) 1 (4) 0 1000 0 644 0 680 - - 17 (2) B, M-C12 DPO (SO ₃ Na) 2 (5) 0 0001 0,000 0,000 - - 18 (2) B, M-C12 DPO (SO ₃ Na) 2 (5) 0 0005 0,000 0.000 - - 9 (2) B, MC, 2 DPO (SO ₃ Na) 2 (5) 0 0010 0,000 0,000 - - 20 (2) B, M-C12 DPO (SO ₃ Na) 2 (5) 0 0100 0 009 0 011 0017 0 005 21 (2) BM-C, 2 DPO (SO ₃ Na) 2 (5) 0 0500 0 027 - - - 22 (2) B, MC, 2 DPO (SO ₃ Na) 2 (5) 0 1000 0 065 0 081 - - 23 L DC 10 DPO (SO ₃ Na), (6) 0 0001 0, 04 - - - 24 L DC 10 DPO (SO ₃ Na), (6) 0 0003 03.0 - - - 25 L DC 10 DPO (SO ₃ Na), (6) 0 0005 0 563 - - - 26 L DC 10 DPO (SO ₃ Na) 1 (6) 0 0010 0 869 - - - 27 L D-C10 DPO (SO ₃ Na), (6) 0 0100 - 0 773 0.605 - 28 L DC 10 DPO (SO ₃ Na) 1 (6) 0 0200 - 0 956 - - 29 (2) L D-C10 DPO (SO ₃ Na) ₃ (7) 0 0001 0 030 - - - 30 (2) L DC 10 DPO (SO ₃ Na) 2 (7) 0 0003 0 041 - - - 31 (2) L DC 10 DPO (SO ₃ Na) ₃ (7) 0 0005 0 096 - - - 32 (2) L DC 10 DPO (SO ₃ Na) 2 (7) 00010 0 164 - - - 33 (2) L DC, DPO (SO ₃ Na) 2 (7) 00100 - 0 444 0 248 - 34 (2) LD-C 10 DPO (SO ₃ Na) ₃ (7) 0 0200 0 581 - - 35 L MC, DPO (SO ₃ Na), (8) 0 0005 0.051 - - - 36 L MC 10 DPO (SO ₃ Na), (8) 0 0010 0, 20 - - - 37 L MC, DPO (SO ₃ Na), (8) 0 0015 0 559 - - - 38 L M-C10 DPO (SO ₃ Na), (8) 00100 - 0 235 0 267 - 39 (2) L M-C10 DPO (SO ₃ Na) ₃ (9) 0 0005 0 0.1 - - - 40 (2) L MC 10 DPO (SO ₃ Na) 2 (9) 0 0010 021 - - - 41 (2) L MC 10 DPO (SO ₃ Na) 2 (9) 0 0015 0 041 - - - 42 (2) L MC, 0 DPO (SO ₃ Na) ₂ (9) 0 0100 - 0.005 0.005 - 43 L DC, DPO (SO ₃ Na) 1 (10) 0 0100 0 744 - 0 889 - 44 (2) L DC DPO (SO ₃ Na) 1 (11) 0.0100 0 289 - 0.522 - 45 L MC, e DPO (SO ₃ Na) 1 (12) 0.0100 0 185 - 0 348 - 46 (2) L MC, e DPO (SO ₃ Na) 2 (13) 0 0100 0 109 - 0 176 - 47 L M-C12 DPO (SO ₃ Na), (14) 00100 - - 0.733 - 48 (2) L MC, 2 DPO (SO ₃ Na) 2 (15) 00100 - - 0.337 -

(1) C12 branched alkyl dialkyl sodium diphenyloximonsulfonate, (2) not according to the invention (3) C12 branched alkyl dialkyloxydisulfonate sodium (4) C12 branched alkyl diphenyloximonsulfonate monoalkylated monoalkylated C12alkylated monoalkylated sodium salt (5) sodium salt Sodium diphenyloximonsulfonate dialkylated with C10 linear alkyl (7) Sodium diphenyloxydisulfonate dialkylated with C10 linear alkyl (8) Sodium diphenyloximonsulfonate monoalkylated with C10 linear alkyl (9) Sodium diphenyloxydisulfonate sodium monoalkylated dialkylated sodium C10 linear alkyl 11) C16 linear alkyl diphenyloxydisulfonate sodium salt (12) C16 linear alkyl monoalkylated diphenyloxymonosulfonate (13) C16 linear alkyl diphenyloxydisulfonate sodium salt (14) C16 linear alkyl diphenyloxymonosulfonate sodium salt C12 linear alkyl dialkylate (15) C12 linear alkyl dialkylated sodium diphenyloxydisulfonate

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The results in Table 9 show that the use of monosulfonated collectors in the process of the invention consistently leads to higher yields of various minerals compared to similar non-sulfonated collectors.

Example X. Sequential flotation.

In this example, the Hallimonda pipe was flotated as described in Example 1. In each case, the feed was a 50/50% by weight mixture of the ingredients listed in Table 10, which also shows the specific collectors used and the mineral yields obtained.

Table 10

A mixture of minerals Mineral yield Collector Dose Ingredient Ingredient kg / kg 1 2 1 2 1 2 3 4 5 6 L DC, 0-DPO (So ₃ Na) 1 (1) 0 025 Apatite Hematite 0 614 0 068 L D-C10-DPO (SOaNa), (1) 0 100 Apatite Hematite 0 947 0 489 L DC, 0-DPO (SO ₃ Na) 1 (1) 0 100 Apatite Dolomite 0 726 0 182 L D-C10-DPO (SOsNa) 1 (1) 0 100 Apatite Dolomite 0.998 0 670 B D-C12-DPO (SOsNa) 1 (2) 0 025 Apatite Martyt 0 873 0 097 B D-C12-DPO (SOaNa), (2) 0 1100 Apatite Martyt 0 944 0 335 B D-C12-DPO (SO ₃ Na) 1 (2) 0 025 Apatite Bauxite 0 604 0 367 B D-C12-DPO (SOjNa), (2) 0 100 Apatite Bauxite 0 889 0.603 BD-C12-DPO (SO ₃ Na), (2) 0 025 Rutile Martyt 0.893 0 223 B D-C, 2-DPO (SOaNa), (2) 0 100 Rutile Martyt 0 947 0.366 B D-C, 2-DPO (SOaNa), (2) 0 025 Rutile Bauxite 0 801 0 229 B D-C12-DPO (SOaNa), (2) 0 100 Rutile Bauxite 0 914 0 377 B D-C12-DPO (SOjNa), (2) 0 025 Gibsite Bemit 0.881 0.137 B D-C12-DPO (SOaNa), (2) 0 100 Gibsite Bemit 0 947 0 229 B D-C10-DPO (SOaNa), (1) 0 025 Gibsite Bemit 0 850 0.111 B D-C10-DPO (SOjNa), (1) 0 100 Gibsite Bemit 0 894 0 203 B D-C, 0-DPO (SOaNa), (1) 0 025 Pyrolusite Hematite 0 717 0 188 B DC, 0-DPO (So ₃ Na) 1 (1) 0.100 Pyrolusite Hematite 0.915 0 404 B D-C12-DPO (SOsNa), (2) 0.025 Topaz Cassiterite 0 791 0 103 B D-C12-DPO (SOaNa), (2) 0.100 Topaz Cassiterite 0.956 0.458 B D-C12-DPO (SOoNa), (2) 0 025 Rutile Kaolin 0.611 0.309 B D-C12-DPO (Sat ₃ Na) 1 (2) 0 100 Rutile Kaolin 0 804 0.518

(1) - Sodium salt of diphenyloximonosulfomate dialkylated with linear C10 alkyl (2) - Sodium salt of diphenyloximonosulfomate dialkylated with branched alkyl C12.

The data presented in the table above show that it is possible to separate various minerals in the flotation process according to the invention by adjusting the collector dose. For example, apatite and dolomite can float together in the method of the invention, but it is obvious that apatite flotates more easily than dolomite at lower collector rates. Thus, the apatite can float in the first step with a lower collector dose, and then the dolomite flotation can be performed with a higher dose. The examination of other tests shows that a similar separation can also be performed with other minerals.

Example XI. Separation of apatite from silica and dolomite.

The procedure described in Example 4 was carried out with the difference that the samples contained 30% apatite, 60% silica and 10% dolomite. In addition, in Runs 2 and 3, purified hydrocarbon was added. The results obtained are presented in Table 11.

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Table 11

Attempt Collector Dose kg / Mg P. Yield P. degree enrichment Mg Yield Mg degree enrichment 1 L D-C10-DPO (SO3Na), (2) 0 050 0 862 0 114 0 391 0 048 2 L D-C, 0-DPO (SO 3Na) 1 (2) 0 050 Refined kerosene (3) 0 050 0 827 0 125 0 320 0 042 3 B D-C, 2-DPO (SO3Na) 2 (4) 0 050 Refined kerosene (3) 0 010 0817 0.135 0.302 0 040 4 (1) Olefinic acid 0 050 Refined kerosene (3) 0 010 0 778 0 107 0 563 0.061

(1) - me according to the invention (2) - Sodium diphenyloximonosulfomate dialkylated with linear C10 alkyl (3) - Refined kerosene - product commercially available from Philips Petroleum as Soltrol kerosene.

The data presented in the table above show that in the presence of the collector used in the process of the invention, apatite flotates better than dolomite and confirms the suitability of the collector for separating apatite and dolomite. Using a standard industrial collector (course 4 in Table 11), a comparable apatite-dolomite separation is not obtained, because phosphorus is significantly contaminated with magnesium. The addition of odor nodes in the process of the invention results in a slight decrease in the phosphorus yield with a higher degree of enrichment with a reduced amount of magnesium collected.

Example XII. Apatite flotation.

In the process of Example 11, a mixture of 30% apatite, 10% calcite and 60% silica was flotated. The obtained results are presented in Table 12.

Table 12

Attempt Collector Dose. kg / Mg P. Yield Degree enrichment 1 L, D-C10DPO (SO3Na) 1 (1) 0.050 0.317 0.128 2 L, D-C10DPO (SO ₃ Na) 1 (1) 0.100 0.792 0.137 3 (2) Oleic acid 0.100 0.551 0.064

(1) - Sodium diphenyloximonosulfomate dilakylated with linear C 10 alkyl (2) - solution not according to the invention

Example XIII. Carbon-based paint flotation.

Five suspensions were prepared, in each case by pulping 240 g of printed paper (70% by weight newsprint and 30% of magazine paper), 1.61 g of diethylenetriaminepentaacetic acid as a color control agent, 10.65 g of sodium silicate, collector in an amount specified in Table 13 and 5.64 g of hydrogen peroxide in water, sufficient to produce a slurry of 2% by weight solids. The pH of the suspensions was 10.5 with the exceptions indicated in the table and the temperature was 45 ° C. The digestion was carried out for 30 minutes. Each slurry was prepared from copies of exactly the same sides to provide a comparable amount of ink in each of the five slurries.

The slurry was transferred to a 15 L Voitha flotation machine with sufficient dilution water to completely fill the flotation cell. Sufficient calcium chloride was added to the pulp to produce water with a hardness of 180 ppm CaCO3. Flotation was initiated by introducing air bubbles and passing through the vigorously mixed pulp for 10 min. The foam was then removed with standard hand paddles to give a flotation product.

The obtained product was filtered and dried. The contents of the flotation cell, which contained cellulose fibers, were also filtered and dried. The flotation product was analyzed

Using a scale of 0-10, where 0 is completely white and 10 is completely black. Cellulose wet mats obtained from the contents of the chambers were examined under a high magnification microscope and the remaining paint particles were counted per unit area.

The results obtained are presented in Table 13. The conditions in all trials are identical with the exceptions indicated in the table.

Table 13

Attempt Collector Dose (g) PH flotation Ink concentration Reading on scale The concentration of the paint visually Assessment of the cellulose mat 1 (1) Oleic acid 55 10 5 4 light gray - 2 L, DC ₁₀ -DPO (SO ₃ Na) 1 (2) 2.0 10 5 5 Gray no change 3 L, D-C10-DPO (SO3Na) 1 (2) twenty 8 0 (3) 6 dark gray decrease by 25% 4 (4) L, DC, 0-DPO (SO ₃ Na) 1 (2) twenty 10.5 8 very dark Gray 50% decrease 5 (4) L, D-C10-DPO (SO ₃ Na) 1 (2) twenty 8 0 (3) 9 light black decrease by 75%

(1) - not according to the invention, commonly used in industry, (2) - sodium salt of diphenyloximonosulfonate dialkylated with linear Cm.

(3) - the pH in the flotation cell was lowered by adding 1N HCl, (4) - In this test, no CaCl2 was added to the flotation machine

The results presented in the table above show that the process of the invention is suitable for separating graphite ink and other carbon-based inks from paper in the deinking process of recycled paper. Trials 2-5, compared to Trial 1, which uses a commonly used industrial collector, shows that the collectors used in the process of the invention lead to a higher paint yield with a much lower collector dosage.

PATTERN

Publishing Department of the UP RP. Circulation of 90 copies. Price: PLN 10,000

Claims (12)

Patent claims A method of recovering minerals by froth flotation, which consists in subjecting an aqueous suspension containing particulate minerals to a froth flotation in the presence of a collector and possibly other flotation additives, under such conditions that the enriched minerals flow to the surface, characterized in that an acid is used as the collector an alkyldiaryloxysulfonic acid or salt thereof or a mixture of such acids or salts in which about 20% of the sulfonic acid or its salt is monosulfonated. 2. The method according to p. The collector as claimed in claim 1, wherein the collector is monosulfonic acid or a salt thereof of the formula shown in the figure, wherein each R is independently a saturated or unsaturated alkyl radical optionally substituted, m and n are independently 0, 1 or 2, each M is independently 0, 1 or 2, hydrogen, alkali metal, alkaline earth metal, ammonium or substituted ammonium, and x and y are each 0 or 1 with the proviso that the sum of x and y is equal to 1. 3. The method according to p. The collector of claim 2, wherein R is an alkyl group having 1 to 24 carbon atoms. 4. The method according to p. 3. The process of claim 2 or 3, characterized in that R is an alkyl group containing 6 to 24 carbon atoms. 5. The method according to p. The process of claim 2 or 3, wherein R is an alkyl group containing from 10 to about 16 carbon atoms. 6. The method according to p. A process according to claim 2 or 3, characterized in that R is a linear or branched alkyl group. 7. The method according to p. The method of claim 2, characterized in that the sum of m and n is 2. 8. The method according to p. The method of claim 1, wherein the total collector concentration is at least 0.001 kg / ton and not more than 5.0 kg / ton. 9. The method according to p. The process of claim 1, wherein the collector is used in which at least 25% of the sulfonic acid or its salts are monosulfonated. 10. The method according to p. The process of claim 9, wherein the collector is at least 40% of the sulfonic acid or its salts monosulfonated. 11. The method according to p. The process of claim 10, wherein the collector is at least 50% monosulfonated of the sulfonic acid or its salts. 12. The method according to p. The process of claim 1, wherein the recovered mineral is graphite and the aqueous slurry comprises paper pulp.