Method for recovering metals in multi-metal sulfide ore associated with uranium, zinc and molybdenum

Abstract

The invention belongs to the field of mineral resource recovery, and particularly relates to a method for recovering metals in multi-metal sulfide ores associated with uranium, zinc and molybdenum, the multi-metal sulfide ores associated with uranium, zinc and molybdenum are subjected to flotation to obtain concentrates and tailings, and the flotation stage comprises the process of carrying out flotation by adopting flotation reagents comprising an inhibitor, a composite collector and a foaming agent; wherein the composite collecting agent comprises a compound # imgabs0 # shown in the formula 1 and a compound # imgabs1 # shown in the formula 2, concentrate and acid liquor are subjected to normal-pressure pre-acid leaching treatment, then the treated concentrate and acid liquor are introduced into an oxygen-containing atmosphere for pressurized oxygen leaching, and pressurized leaching ore pulp is obtained; the pressure leaching ore pulp, the tailings and an oxidizing agent are subjected to normal-pressure oxidation leaching, then solid-liquid separation is conducted, and leaching liquid enriched with uranium, zinc and molybdenum is obtained; and uranium, zinc and molybdenum are separated and recovered from the leachate. The mineral physicochemical characteristic limitation of the polymetallic sulfide ore associated with uranium, zinc and molybdenum can be broken through, uranium, zinc and molybdenum in the polymetallic sulfide ore can be efficiently recycled, sulfur in the ore can be efficiently and reasonably utilized, the dosage of sulfuric acid added in the treatment process can be remarkably reduced, and the cost is remarkably reduced.

Description

Technical Field

[0001] The invention belongs to the technical field of hydrometallurgy, and in particular relates to a method for comprehensively recovering metals from polymetallic sulfide ores associated with uranium, zinc and molybdenum. Background Art

[0002] Polymetallic sulfide deposits with associated uranium, zinc and molybdenum are a common type of polymetallic combination deposits. Their production is often related to Mesozoic volcanic activity or volcanic rocks. For example, the Hongshanzi mineralization belt in northern my country produces rhyolite from the Zhangjiakou Formation of the Early Cretaceous of the Mesozoic, and its genetic type is a typical volcanic hydrothermal deposit. This type of deposit is mainly concealed, with the ore body generally buried more than 200 meters below the surface, and there are significant mineralization clues and weak mineralization information on the surface. Some deposits in this mineralization belt contain a large number of polymetallic sulfide ores with associated uranium, zinc and molybdenum with low metal grades. There are currently no public reports on the treatment of this type of polymetallic sulfide molybdenum ores with associated uranium and zinc. The use of existing general technologies to treat this type of polymetallic ores has the following shortcomings:

[0003] (1) When using mineral processing technology for enrichment, due to the complex and diverse occurrence of molybdenum, in addition to molybdenum sulfide, some molybdenum is present in different gangue minerals in the form of molybdenum oxide. Therefore, it is difficult to achieve efficient enrichment of molybdenum using mineral processing technology. In addition, the flotation tailings contain metallic uranium and zinc and cannot be directly discarded.

[0004] (2) When using the acid method for direct agitation leaching, on the one hand, the consumption of leaching agent and oxidant is large. On the other hand, since some sulfide minerals are encapsulated, direct agitation leaching is difficult to break the encapsulation of sulfide minerals, which will lead to low leaching rates of molybdenum and zinc.

[0005] (3) When the raw ore is directly treated by oxygen pressure acid leaching, on the one hand, the processing volume of the raw ore is large, the investment in oxygen pressure equipment is large, and the processing cost is high. On the other hand, a large amount of gangue minerals are dissolved and leached during pressurized acid leaching oxidation, resulting in a high concentration of impurity ions in the leachate, high reagent consumption, difficulty in solid-liquid separation of the ore slurry, and complex subsequent solution treatment process.

[0006] (4) When using the roasting pretreatment method, on the one hand, the equipment investment is large and the treatment cost is high. On the other hand, due to the restrictions of environmental protection requirements, some places restrict the application of pyro-roasting technology.

[0007] In summary, the existing technology does not have a recovery method that can effectively adapt to the characteristics of polymetallic sulfide mineralization with associated uranium, zinc and molybdenum. Summary of the invention

[0008] The purpose of the present invention is to provide a set of methods suitable for the comprehensive treatment of complex low-grade polymetallic sulfide deposits with associated uranium, zinc and molybdenum, so as to achieve comprehensive and efficient recovery of uranium, zinc and molybdenum resources.

[0009] Polymetallic sulfide ores with associated uranium, zinc and molybdenum, wherein the uranium grade in the polymetallic ore is generally 0.03% to 0.10%, the zinc grade is generally 0.5% to 5%, and the molybdenum grade is generally 0.5% to 2%. Part of the uranium exists in an adsorbed state, and part of it is filled in small cracks together with pyrite and zinc sulfide as independent minerals such as pitchblende and uraninite to form fine veins. Molybdenum mainly exists in the form of molybdenum calcium ore, uranium-molybdenum ore, blue molybdenum ore and molybdenum sulfide ore, and part of it exists in the form of molybdenum oxide in gangue minerals. The associated uranium, zinc and molybdenum in this type of minerals interfere with each other, making it difficult to effectively recover them. Aiming at the problem of difficulty in recycling uranium, zinc and molybdenum faced by this special mineral, the present invention provides the following solutions after in-depth research:

[0010] A method for recovering metals from polymetallic sulfide ores associated with uranium, zinc and molybdenum, comprising the following steps:

[0011] Step (1):

[0012] The polymetallic sulfide ores with associated uranium, zinc and molybdenum are subjected to flotation to obtain concentrates and tailings, wherein the flotation stage includes a flotation process using a flotation agent including an inhibitor, a composite collector and a frother; wherein the composite collector includes a compound of formula 1 and a compound of formula 2;

[0013]

[0014] The R1, R2, and R3 are independently C2-C8 alkyl groups; M and N are independently H, Na, K, or NH4;

[0015] Step (2):

[0016] The concentrate and the acid solution are subjected to normal pressure pre-acid leaching treatment and then introduced into an oxygen-containing atmosphere for pressurized oxygen leaching to obtain a pressurized leached slurry;

[0017] Step (3):

[0018] The pressure leaching slurry, the tailings and the oxidant are subjected to normal pressure oxidation leaching, followed by solid-liquid separation to obtain a leaching solution enriched with uranium, zinc and molybdenum; and uranium, zinc and molybdenum are separated and recovered from the leaching solution.

[0019] In view of the problem that uranium, zinc and molybdenum are difficult to be effectively recovered due to the special physical and chemical characteristics of polymetallic sulfide ores with associated uranium, zinc and molybdenum, the present invention innovatively adopts the flotation agent to carry out flotation treatment on the mineral, and then cooperates with subsequent pressurized oxygen leaching and a tailings atmospheric pressure oxygen leaching process involving a leaching system based on pressurized oxygen leaching, so as to break through the mineral physical and chemical characteristics limitations of the polymetallic sulfide ores with associated uranium, zinc and molybdenum, and efficiently recover uranium, zinc and molybdenum therein. Moreover, the process described in the present invention can efficiently and reasonably utilize sulfur in the mineral, so as to significantly reduce the amount of sulfuric acid added in the treatment process and significantly reduce costs.

[0020] In the present invention, the combined control of the flotation and the flotation reagents is the key to solve the physical and chemical limitations of the special minerals, and then efficiently utilize the sulfur in the minerals and effectively improve the recovery rates of uranium, zinc and molybdenum, by synergistic combination with the subsequent special pressurized oxygen leaching and tailings auxiliary atmospheric pressure oxygen leaching involving the pressurized oxygen leaching system.

[0021] In the present invention, the combined flotation of the inhibitor and the composite collector can optimize the flotation of uranium, zinc and molybdenum in the raw ore, which is beneficial to the combination with the subsequent pressurized oxygen leaching and normal pressure oxygen leaching to strengthen the recovery of uranium, zinc and molybdenum in the special minerals.

[0022] In the present invention, the inhibitor comprises at least one of water glass, sodium fluorosilicate and sodium hexametaphosphate.

[0023] In the present invention, in the composite collector, the weight ratio of the compound of formula 1 to the compound of formula 2 is 10-50:5-30.

[0024] In the present invention, the foaming agent includes at least one of No. 2 oil, rosin oil, methyl isobutyl alcohol (MIBC) and polyethylene glycol.

[0025] In the present invention, the flotation procedure can be conventional. For example, an optional flotation process of the present invention includes a first roughing process and two-stage scavenging processes, and the steps are: crushing and grinding the polymetallic sulfide ore to obtain ore pulp, and then compounding it with the flotation agent for a first roughing treatment to obtain a roughing concentrate and a roughing tailing; using a flotation agent a containing a composite collector and an inhibitor to perform a first-stage scavenging treatment on the roughing tailings to obtain a first-stage scavenging concentrate and a first-stage scavenging tailings; then using a flotation agent b containing the composite collector to perform a second-stage scavenging treatment on the first-stage scavenging tailings to obtain a second-stage scavenging concentrate and a second-stage scavenging tailings; the second-stage scavenging concentrate is returned to the first-stage scavenging process, and the first-stage scavenging concentrate is returned to the roughing process; the second-stage scavenging tailings are the tailings of flotation; and the roughing concentrate is the concentrate of flotation.

[0026] In the present invention, the operations of crushing, grinding, flotation, etc. can be achieved based on conventional means.

[0027] In the present invention, the reagents in the flotation stage can be adjusted as needed.

[0028] In the roughing process, the dosage of the inhibitor is 500-5000 g / t, the dosage of the composite collector is 150-800 g / t, and the dosage of the foaming agent is 50-200 g / t. Further, the dosage of the inhibitor is 2000-3500 g / t, and further 2500-3000 g / t; the dosage of the composite collector is 200-400 g / t, and further 250-350 g / t; and the dosage of the foaming agent is 80-100 g / t.

[0029] In the present invention, in the first scavenging process, the amount of the inhibitor is 200-1000 g / t, and the amount of the composite collector is 100-200 g / t. Further, in the first scavenging process, the amount of the inhibitor is 500-1500 g / t, further 800-1000 g / t, and the amount of the composite collector is 150-250 g / t.

[0030] In the present invention, the dosage of the composite collector in the second scavenging process is 50-200 g / t, and further can be 100-150 g / t.

[0031] In the present invention, thanks to the combination of the flotation reagents, further combined with subsequent special acid leaching, it is possible to efficiently recover uranium, zinc and molybdenum in the special minerals.

[0032] In the present invention, in step (2), the acid solution is a sulfuric acid solution, and the concentration of the solute in the acid solution is 0.1 to 0.3M.

[0033] In the present invention, in the atmospheric pressure pre-acid leaching stage, the weight of the solute in the acid solution is 1-5% of the weight of the concentrate.

[0034] In the present invention, in the normal pressure pre-acid leaching stage, the liquid-to-solid ratio of the acid solution to the concentrate is 1-3 mL / g.

[0035] In the present invention, in step (2), the oxygen-containing atmosphere is oxygen or air.

[0036] In the present invention, the reaction pressure in the pressurized oxygen immersion stage is 1.5-2.0 MPa, the reaction temperature is 180-200° C., the reaction oxygen partial pressure is 0.5-0.8 MPa, and the reaction time is 3-6 hours.

[0037] In the present invention, in step (3), the concentration of sulfuric acid in the pressure leaching slurry at the beginning of atmospheric pressure oxygen leaching is controlled to be 10-20 g / L.

[0038] In the present invention, the liquid-to-solid ratio of the pressure leaching slurry and tailings in the atmospheric pressure oxygen leaching stage is 1 to 3 mL / g.

[0039] In the present invention, the oxidant includes at least one of hydrogen peroxide, hypochlorous acid and potassium permanganate.

[0040] In the present invention, the potential of the leaching solution is controlled at 450-500 mV during the atmospheric pressure oxygen leaching stage.

[0041] The present invention is based on the combination of the flotation process and the leaching process, and can efficiently extract uranium, zinc and molybdenum therein. Not only that, it also helps to improve the recovery selectivity of uranium, zinc and molybdenum, and thus facilitates the subsequent effective separation of uranium, zinc and molybdenum.

[0042] In step (4) of the present invention, the leaching solution is preliminarily subjected to co-adsorption of uranium and molybdenum by anion exchange adsorption, and then desorbed step by step to obtain qualified uranium solution and qualified molybdenum solution respectively; and the adsorption tail liquid in the co-adsorption stage is deironized and then subjected to cation exchange adsorption to adsorb zinc, and then desorbed to obtain qualified zinc solution.

[0043] In the present invention, in step (4), the resin in the anion exchange adsorption stage includes at least one of the macroporous anion exchange resins D382, D314, D301 and D816.

[0044] In the present invention, the desorption solution for desorbing uranium comprises a solution of 0.5-1M sodium chloride and 0.01-0.1M sulfuric acid.

[0045] In the present invention, the desorption liquid for desorbing molybdenum is 5-10% ammonia water.

[0046] In the present invention, the resin for cation exchange adsorption includes at least one of D152, D113, and D155.

[0047] In the present invention, the desorption liquid for desorbing zinc is dilute sulfuric acid with a mass concentration of 8 to 15%.

[0048] Beneficial Effects

[0049] The present invention innovatively adopts the flotation reagent to carry out flotation treatment on the mineral, and then cooperates with the subsequent pressurized oxygen leaching and the combination of the leachate based on the pressurized oxygen leaching and the normal pressure oxygen leaching of the tailings, so as to break through the mineral physicochemical characteristics of the polymetallic sulfide ore associated with uranium, zinc and molybdenum, improve the utilization rate of mineral sulfur, and obtain excellent uranium, zinc and molybdenum recovery rate and recovery selectivity at low acid dosage. For example, the amount of sulfuric acid used in the process of the present invention can be controlled within 1.5-3%.

[0050] The present invention is based on the co-adsorption of U and Mo by the anion resin and the subsequent adsorption of Zn by the cation resin, and further cooperates with the combination of desorption means to simply obtain U, Mo and Zn products. BRIEF DESCRIPTION OF THE DRAWINGS

[0051] Figure 1 It is a schematic diagram of the process flow of the present invention; DETAILED DESCRIPTION

[0052] The present invention is further described below with reference to the embodiments, but the contents of the present invention are not limited to the described scope.

[0053] The present invention discloses a process for recovering polymetallic sulfide ores associated with uranium, zinc and molybdenum, the steps of which include:

[0054] (1) Ore fine grinding and slurry preparation: The polymetallic sulfide ores with associated uranium, zinc and molybdenum are crushed and finely ground before slurry preparation.

[0055] (2) Sulfide flotation enrichment: Using water glass as the inhibitor, Formula 1 and Formula 2 as the collector, and No. 2 oil as the frother, after one roughing + two sweeping processes, flotation concentrate and flotation tailings are obtained.

[0056] (3) Oxygen pressure acid leaching of flotation concentrate: After the flotation sulfide mineral concentrate is pre-leached with sulfuric acid, oxygen is introduced for pressurized oxygen leaching to obtain a high-acidity leaching slurry.

[0057] (4) Acid leaching of flotation tailings at normal pressure: The high-acid leaching pulp from step (3), the flotation tailings and the oxidant are mixed and subjected to oxidative acid leaching at normal pressure to obtain a multi-metal mixed leaching solution after solid-liquid separation.

[0058] (5) Metal separation of mixed leaching solution: First, uranium and molybdenum are co-adsorbed by anion exchange adsorption, and qualified uranium solution and molybdenum solution are obtained after step-by-step desorption. The pH value of the adsorption tail liquid is adjusted to remove iron, and then zinc is adsorbed by cation exchange adsorption. After desorption, qualified zinc solution is obtained.

[0059] The step (1) specifically comprises: crushing and grinding the ore to a particle size of less than 0.074 mm, with a mass percentage content of 50% to 90%, and adjusting the mass percentage concentration of the ore pulp to 20% to 40%.

[0060] The step (2) is specifically as follows: in the roughing process, the amount of water glass is 500-5000 g / t, the amount of formula 1 is 100-500 g / t, the amount of formula 2 is 50-300 g / t, and the amount of No. 2 oil is 50-200 g / t; in the primary scavenging process, the amount of water glass is 200-1000 g / t, the amount of formula 1 is 50-100 g / t, and the amount of formula 2 is 50-100 g / t; in the secondary scavenging process, the amount of formula 1 is 25-75 g / t, and the amount of formula 2 is 25-75 g / t.

[0061] The step (3) is specifically as follows: adding 1% to 5% sulfuric acid by mass percentage of the flotation concentrate obtained in step (2) for pre-leaching (the pre-leaching time can be more than 0.2h, and can further be 0.5 to 1h), and adding process water to make the solid volume mass ratio of the pre-leached slurry be 1 to 3mL / g, introducing oxygen after acid pre-leaching, controlling the reaction pressure to 1.5 to 2.0MPa, the reaction temperature to 180 to 200°C, the reaction oxygen partial pressure to 0.5 to 0.8MPa, and the reaction time to 3 to 6h to obtain the concentrate oxygen pressure leaching slurry.

[0062] The step (4) is specifically as follows: the high-acidity leaching pulp obtained in step (3) is mixed with the flotation tailings for acid leaching at normal pressure, the amount of sulfuric acid added is preferably such that the residual sulfuric acid mass concentration of the leaching solution is controlled to be 10-20 g / L, and process water is added to make the slurry liquid-solid volume mass ratio be 1-3 mL / g, and the added oxidant is one of hydrogen peroxide, hypochlorous acid, and potassium permanganate, and the added amount is such that the leaching solution potential is controlled to be 450-500 mV. After solid-liquid separation, a multi-metal mixed leaching solution is obtained.

[0063] The step (5) specifically comprises: firstly using a plurality of columns in series to synchronously adsorb uranium and molybdenum on the polymetallic mixed leaching solution obtained in the step (4), wherein the resin is one of the macroporous anion exchange resins D382, D314, D301 and D816, and the adsorption contact time is 10 to 30 minutes; after the resin is saturated with adsorption, using a 0.6 to 1 mol / L NaCl + 0.03 to 0.08 mol / L H2SO4 solution as a desorbent for uranium, and the desorption volume is 9 to 14 times the resin bed volume, and then using ammonia water with a mass concentration of 5 to 10% as a desorbent for molybdenum, and the desorption contact time is 20 to 40 minutes, and the desorption volume is 6 to 10 times the resin bed volume. The uranium-molybdenum adsorption tail liquid is adjusted to pH 3.5-5 by adding alkali, and after precipitation and iron removal, zinc is recovered by adsorption with a cation exchange resin, the resin is one of D152, D113, and D155, the adsorption contact time is 10-30 minutes, and the adsorption saturated resin column is desorbed with dilute sulfuric acid with a mass concentration of 8-15%, the desorption contact time is 20-40 minutes, and the desorption volume is 2-4 times the volume of the resin bed. Through the step-by-step adsorption and desorption of the two resins, uranium qualified liquid, molybdenum qualified liquid and zinc qualified liquid are obtained respectively.

[0064] In the following cases, Formula 1 is represented by Formula 1a as a typical scheme, wherein Formula 1a is a compound of Formula 1 in which R1 is n-butyl and M is Na.

[0065] The formula 2 is represented by formula 2a as a typical scheme, wherein the formula 2a is a compound of formula 2 in which R2 and R3 are both n-butyl and N is NH4.

[0066] In the following cases, the unit of dosage of the flotation reagent is g / t, that is, the gram weight (g) of the reagent required per ton of mineral to be floated.

[0067] Example 1

[0068] A polymetallic sulfide ore with associated uranium, zinc and molybdenum in Hebei has a uranium grade of 0.05%, a zinc grade of 2.80%, a molybdenum grade of 1.20% and a sulfur content of 8%. Part of the uranium exists in an adsorbed state, and part of it is filled in small cracks with pyrite and zinc sulfide as independent minerals such as pitchblende and uranium stone to form fine veins. Molybdenum mainly exists in the form of molybdenum calcium ore, uranium-molybdenum ore, blue molybdenum ore and molybdenum sulfide ore, and part of it exists in the form of molybdenum oxide in gangue minerals. The treatment method described in the present invention includes the following process steps:

[0069] (1) Grind the ore to a particle size of -0.074 mm, accounting for 70%, and adjust the slurry concentration to 20%.

[0070] (2) 3000 g / t of water glass was added to the slurry, and the mixture was stirred for 10 min at a flotation machine rotor speed of 1500 r / min; then 200 g / t of formula 1a and 100 g / t of formula 2a were added, and the mixture was stirred for 10 min at a flotation machine rotor speed of 1500 r / min; then 50 g / t of No. 2 oil was added, and the mixture was stirred for 10 min at a flotation machine rotor speed of 1500 r / min to obtain a roughing concentrate and a roughing tailing. In the rougher tailings, 1000 g / t of water glass, 150 g / t of formula 1a, and 75 g / t of formula 2a were added, and the mixture was stirred for 10 min at a flotation machine rotor speed of 1500 r / min to obtain scavenger concentrate 1 and scavenger tailing 1; 100 g / t of formula 1a and 50 g / t of formula 2a were added to the scavenger tailings 1, and the mixture was stirred for 10 min at a flotation machine rotor speed of 1500 r / min to obtain scavenger concentrate 2 and scavenger tailing 2. The rougher concentrate and the secondary scavenger concentrate were combined as the flotation concentrate, and the scavenger tailing 2 was the obtained flotation tailing.

[0071] Table 1 Flotation results of Example 1

[0072]

[0073]

[0074] (3) Pre-leaching the flotation concentrate obtained in step (2) by adding 3% sulfuric acid by mass of the concentrate, and adding process water so that the solid volume mass ratio of the pre-leached slurry is 1.5 mL / g. After acid pre-leaching for 0.5 to 1 h, oxygen is introduced, and the reaction pressure is controlled to be 1.6 MPa, the reaction temperature is 180° C., the reaction oxygen partial pressure is 0.6 MPa, and the reaction time is 5 h to obtain a concentrate pressure leaching slurry with a sulfuric acid mass concentration of 66.8 g / L.

[0075] (4) The pressurized leaching slurry obtained in step (3) is mixed with the flotation tailings for normal pressure acid leaching, the amount of sulfuric acid added is 2% of the mass percentage of the tailings, and process water is added to make the slurry liquid-solid volume mass ratio 2mL / g, the added oxidant is 27.5% hydrogen peroxide, the addition amount is 20mL / L, and after solid-liquid separation, tailings and multi-metal mixed leaching solution are obtained, the residual sulfuric acid mass concentration of the leaching solution is 18.2g / L, and the potential is 478mV.

[0076] (5) The polymetallic mixed leaching solution obtained in step (4) is firstly connected in series with three columns to synchronously adsorb uranium and molybdenum, and the resin is a macroporous anion exchange resin D301, and the adsorption contact time is 15 minutes to obtain a UMo-loaded resin and an adsorption tail liquid; after the first column resin is saturated with adsorption, 0.8 mol / L NaCl + 0.05 mol / L H2SO4 solution is used as a desorbent for uranium, and the desorption volume is 9 times the resin bed volume, and then ammonia water with a mass concentration of 6% is used as a desorbent for molybdenum, and the desorption contact time is 30 minutes, and the desorption volume is 6 times the resin bed volume.

[0077] The above uranium-molybdenum adsorption tail liquid was adjusted to pH 4 by adding calcium hydroxide, and after precipitation to remove trivalent iron, zinc was adsorbed and recovered by D152 cation exchange resin, and the adsorption contact time was 20 minutes. The adsorption saturated resin column was desorbed by dilute sulfuric acid with a mass concentration of 10%, and the desorption contact time was 30 minutes. The desorption volume was 3 times the resin bed volume. The concentrations of 11.5g / L uranium qualified solution, 55.5g / L molybdenum qualified solution, and 72.5g / L zinc qualified solution were obtained respectively.

[0078] Final leaching effect: tailings uranium grade 0.008%, molybdenum grade 0.158%, zinc grade 0.758%; slag uranium leaching rate 84.01%, molybdenum leaching rate 86.84%, zinc leaching rate 72.94%; total sulfuric acid dosage 2.78% (referring to the percentage of sulfuric acid added during the whole process relative to the weight of the mineral, the same below).

[0079] Example 2

[0080] A polymetallic sulfide ore with associated uranium, zinc and molybdenum is mainly composed of quartz, plagioclase, microcline, dolomite, calcite, gypsum, pitchblende, pyrite, sphalerite, goethite, limonite, molybdenum sulfide, etc., and the ore has a high sulfide content. Uranium exists in two forms: independent mineral and adsorption state. Uranium minerals include pitchblende, uraninite, and titanouraninite. The output of most uranium minerals is closely related to pyrite, sphalerite, feldspar and quartz. Molybdenum mainly exists in the form of uranium-molybdenum ore and molybdenum sulfide ore, and a part of it exists in the form of molybdenum oxide in gangue minerals. The uranium grade in the ore is 0.07%, the zinc grade is 3.22%, the molybdenum grade is 0.959%, ​​and the sulfur content is 7.5%. The treatment method described in the present invention includes the following process steps:

[0081] (1) Grind the ore to a particle size of -0.074 mm, accounting for 75%, and adjust the slurry concentration to 30%.

[0082] (2) Add 2500 g / t of water glass to the ore pulp, and stir for 12 min at a flotation machine rotor speed of 2000 r / min; then add 180 g / t of formula 1a and 80 g / t of formula 2a, and stir for 15 min at a flotation machine rotor speed of 2000 r / min; then add 100 g / t of No. 2 oil, and stir for 12 min at a flotation machine rotor speed of 2000 r / min to obtain a roughing concentrate and a roughing tailing. In the rougher tailings, 800 g / t of water glass, 120 g / t of formula 1a, and 50 g / t of formula 2a were added, and the mixture was stirred for 12 min at a flotation machine rotor speed of 2000 r / min to obtain scavenger concentrate 1 and scavenger tailing 1; 80 g / t of formula 1a and 30 g / t of formula 2a were added to the scavenger tailings 1, and the mixture was stirred for 12 min at a flotation machine rotor speed of 2000 r / min to obtain scavenger concentrate 2 and scavenger tailing 2. The rougher concentrate and the secondary scavenger concentrate were combined as the flotation concentrate, and the scavenger tailing 2 was the obtained flotation tailing.

[0083] Table 2 Flotation results of Example 2

[0084]

[0085] (3) Pre-leaching the flotation concentrate obtained in step (2) by adding sulfuric acid with a mass percentage of 2.5% of the concentrate, and adding process water so that the solid volume mass ratio of the pre-leached slurry is 1.8 mL / g. After acid pre-leaching for 0.5 to 1 h, oxygen is introduced, and the reaction pressure is controlled to be 1.65 MPa, the reaction temperature is 190° C., the reaction oxygen partial pressure is 0.62 MPa, and the reaction time is 4 h to obtain a concentrate pressure leaching slurry with a sulfuric acid mass concentration of 72.5 g / L.

[0086] (4) The pressurized leaching slurry obtained in step (3) is mixed with the flotation tailings for normal pressure acid leaching, the amount of sulfuric acid added is 2.2% of the mass percentage of the tailings, and process water is added to make the slurry liquid-to-solid mass ratio 2 mL / g, the added oxidant is 27.5% hydrogen peroxide, the amount added is 15 mL / L, and after solid-liquid separation, a multi-metal mixed leaching solution is obtained, the residual sulfuric acid mass concentration of the leaching solution is 15.8 g / L, and the potential is 475 mV.

[0087] (5) The polymetallic mixed leachate obtained in step (4) is firstly connected in series with three columns to synchronously adsorb uranium and molybdenum, and the resin is macroporous anion exchange resin D382, and the adsorption contact time is 20 minutes; after the first column resin is saturated with adsorption, 0.8mol / LNaCl+0.05mol / L H2SO4 solution is used as the desorbent for uranium, and the desorption volume is 10 times the resin bed volume, and then ammonia water with a mass concentration of 6% is used as the desorbent for molybdenum, and the desorption contact time is 25 minutes, and the desorption volume is 7 times the resin bed volume. The above uranium and molybdenum adsorption tail liquid is added with lime milk to adjust the pH to 4.5, and after precipitation to remove trivalent iron, zinc is adsorbed and recovered by D113 cation exchange resin, and the adsorption contact time is 25 minutes. The adsorption saturated resin column is desorbed with dilute sulfuric acid with a mass concentration of 8%, and the desorption contact time is 30 minutes, and the desorption volume is 4 times the resin bed volume. The obtained concentrations were 12.88 g / L uranium qualified solution, 48.4 g / L molybdenum qualified solution, and 83.3 g / L zinc qualified solution.

[0088] The final leaching effect: the tailings uranium grade is 0.008%, the molybdenum grade is 0.165%, and the zinc grade is 0.865%; the slag-based uranium leaching rate is 88.54%, the molybdenum leaching rate is 82.80%, and the zinc leaching rate is 73.17%; the total sulfuric acid consumption is 2.27%.

[0089] Comparative Example 1

[0090] Compared with Example 1, the only difference is that in the flotation stage of step 2, Formula 1a is replaced by Formula 2a of the same amount, and other operations and parameters are the same as in Example 1.

[0091] The flotation result of step 2 is: a flotation concentrate with a yield of 25.65% and an S grade of only 14.52% can be obtained, the sulfur recovery rate is 46.55%, and the total sulfuric acid consumption is increased from 2.78% to 9.85%.

[0092] Comparative Example 2

[0093] Compared with Example 1, the only difference is that in the flotation stage of step 2, Formula 2a is replaced by Formula 1a of the same amount, and other operations and parameters are the same as in Example 1.

[0094] The flotation result of step 2 is: only a flotation concentrate with a yield of 18.85% and an S grade of 15.8% can be obtained, the sulfur recovery rate is 37.22%, and the total sulfuric acid consumption increases from 2.78% to 10.5%.

[0095] Comparative Example 3

[0096] Compared with Example 1, the only difference is that step 2 is omitted, and step 3 and subsequent treatments are directly performed on the mineral in step 1. Other operations and parameters are the same as those in Example 1.

[0097] The results are as follows: the uranium grade of the tailings is 0.010%, the molybdenum grade is 0.188%, and the zinc grade is 0.928%; the uranium leaching rate of the slag is 80.0%, the molybdenum leaching rate is 84.33%, and the zinc leaching rate is 66.86%; and the total sulfuric acid consumption is 5.5%.

[0098] Comparative Example 4

[0099] Compared with Example 1, the only difference is that in step 4, the slurry of step 3 is not used to treat the tailings, but sulfuric acid with equal acidity is used to replace the slurry of step 3 to leach the tailings, and the leaching system of step 3 and the leaching system of step 4 are combined to perform the treatment of step 5, and other operations and parameters are the same as in Example 1.

[0100] The result was that the total sulfuric acid usage increased from 2.78% to 15.8%.

[0101] The present invention discloses a method for comprehensively recovering metals from polymetallic sulfide ores with associated uranium, zinc and molybdenum. The method groups the ores by flotation according to the mineral properties of the ores and the occurrence characteristics of molybdenum, which can effectively reduce the reagent consumption of subsequent leaching and improve the leaching rate of metals. Different leaching and separation and extraction processes are combined to achieve the separation and recovery of uranium, molybdenum and zinc polymetallics, thus providing technical support for the economic development of such polymetallic mineral resources.

Claims

1. A method for recovering metals from polymetallic sulfide ores containing uranium, zinc and molybdenum, characterized in that the steps include: Step (1): The polymetallic sulfide ores with associated uranium, zinc and molybdenum are subjected to flotation to obtain concentrates and tailings, wherein the flotation stage includes a flotation process using a flotation agent including an inhibitor, a composite collector and a frother; wherein the composite collector includes a compound of formula 1 and a compound of formula 2; The R1, R2, and R3 are independently C2-C8 alkyl groups; M and N are independently H, Na, K, or NH4; Step (2): The concentrate and the acid solution are subjected to normal pressure pre-acid leaching treatment and then introduced into an oxygen-containing atmosphere for pressurized oxygen leaching to obtain a pressurized leached slurry; Step (3): The pressure leaching slurry, the tailings and the oxidant are subjected to normal pressure oxidation leaching, followed by solid-liquid separation to obtain a leaching solution enriched with uranium, zinc and molybdenum; and uranium, zinc and molybdenum are separated and recovered from the leaching solution.

2. The method for recovering metals from polymetallic sulfide ores associated with uranium, zinc and molybdenum as claimed in claim 1, characterized in that: In the polymetallic sulfide ore with associated uranium, zinc and molybdenum, the uranium grade is 0.03% to 0.10%, the zinc grade is 0.5% to 5%, and the molybdenum grade is 0.5% to 2%.

3. The method for recovering metals from polymetallic sulfide ores associated with uranium, zinc and molybdenum as claimed in claim 1, characterized in that: The inhibitor comprises at least one of water glass, sodium fluoride silicate and sodium hexametaphosphate; Preferably, in the composite collector, the weight ratio of the compound of formula 1 to the compound of formula 2 is 10-50:5-30; Preferably, the foaming agent includes at least one of No. 2 oil, rosin oil, methyl isobutyl alcohol (MIBC), and polyethylene glycol.

4. The method for recovering metals from polymetallic sulfide ores associated with uranium, zinc and molybdenum as claimed in claim 1 or 3, characterized in that: The flotation process includes a first roughing process and two-stage scavenging processes, and the steps are: crushing and grinding the polymetallic sulfide ore to obtain ore pulp, and then compounding it with the flotation agent for a first roughing treatment to obtain a roughing concentrate and a roughing tailing; using a flotation agent a containing a composite collector and an inhibitor to perform a first-stage scavenging treatment on the roughing tailings to obtain a first-stage scavenging concentrate and a first-stage scavenging tailings; then using a flotation agent b containing the composite collector to perform a second-stage scavenging treatment on the first-stage scavenging tailings to obtain a second-stage scavenging concentrate and a second-stage scavenging tailings; the second-stage scavenging concentrate is returned to the first-stage scavenging process, and the first-stage scavenging concentrate is returned to the roughing process; the second-stage scavenging tailings are the tailings of flotation; and the roughing concentrate is the concentrate of flotation.

5. The method for recovering metals from polymetallic sulfide ores associated with uranium, zinc and molybdenum as claimed in claim 4, characterized in that: In the roughing process, the dosage of the inhibitor is 500-5000 g / t, the dosage of the composite collector is 150-800 g / t, and the dosage of the foaming agent is 50-200 g / t; Preferably, in the first sweeping process, the dosage of the inhibitor is 200-1000 g / t, and the dosage of the composite collector is 100-200 g / t; Preferably, the dosage of the composite collector in the second scavenging process is 50-200 g / t.

6. The method for recovering metals from polymetallic sulfide ores associated with uranium, zinc and molybdenum as claimed in claim 1, characterized in that: In step (2), the acid solution is a sulfuric acid solution; preferably, the concentration of the solute in the acid solution is 0.1 to 0.3 M; Preferably, in the atmospheric pressure pre-acid leaching stage, the weight of the solute in the acid solution is 1 to 5% of the weight of the concentrate; Preferably, in the atmospheric pressure pre-acid leaching stage, the liquid-to-solid ratio of the acid solution to the concentrate is 1 to 3 mL / g.

7. The method for recovering metals from polymetallic sulfide ores associated with uranium, zinc and molybdenum as claimed in claim 1, characterized in that: In step (2), the oxygen-containing atmosphere is oxygen or air; Preferably, the reaction pressure in the pressurized oxygen immersion stage is 1.5-2.0 MPa, the reaction temperature is 180-200° C., the reaction oxygen partial pressure is 0.5-0.8 MPa, and the reaction time is 3-6 h.

8. The method for recovering metals from polymetallic sulfide ores associated with uranium, zinc and molybdenum as claimed in claim 1, characterized in that: In step (3), the concentration of sulfuric acid in the pressure leaching slurry at the beginning of atmospheric pressure oxygen leaching is controlled to be 10 to 20 g / L; Preferably, the liquid-to-solid ratio of the pressure leaching slurry and tailings in the atmospheric pressure oxygen leaching stage is 1 to 3 mL / g; Preferably, the oxidant comprises at least one of hydrogen peroxide, hypochlorous acid, and potassium permanganate; Preferably, the potential of the leaching solution is controlled at 450-500 mV during the atmospheric pressure oxygen leaching stage.

9. The method for recovering metals from polymetallic sulfide ores associated with uranium, zinc and molybdenum as claimed in claim 1, characterized in that: In step (4), the leaching solution is preliminarily subjected to co-adsorption of uranium and molybdenum by anion exchange adsorption, and then desorbed step by step to obtain qualified uranium solution and qualified molybdenum solution respectively; and the adsorption tail liquid in the co-adsorption stage is deironized and then subjected to cation exchange adsorption to adsorb zinc, and then desorbed to obtain qualified zinc solution.

10. The method for recovering metals from polymetallic sulfide ores associated with uranium, zinc and molybdenum as claimed in claim 9, characterized in that: In step (4), the resin in the anion exchange adsorption stage includes at least one of macroporous anion exchange resins D382, D314, D301 and D816; Preferably, the desorption solution for desorbing uranium includes 0.5-1M sodium chloride and 0.01-0.1M sulfuric acid solution; Preferably, the desorption liquid for desorbing molybdenum is 5-10% ammonia water; Preferably, the cation exchange adsorption resin includes at least one of D152, D113, and D155; Preferably, the desorption solution for desorbing zinc is dilute sulfuric acid with a mass concentration of 8 to 15%.

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