WO2000048944A1 - Improved methods for leaching of ores - Google Patents

Improved methods for leaching of ores Download PDF

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Publication number
WO2000048944A1
WO2000048944A1 PCT/US2000/004333 US0004333W WO0048944A1 WO 2000048944 A1 WO2000048944 A1 WO 2000048944A1 US 0004333 W US0004333 W US 0004333W WO 0048944 A1 WO0048944 A1 WO 0048944A1
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WO
WIPO (PCT)
Prior art keywords
alkali metal
acid
halide
solubihzation
titanium
Prior art date
Application number
PCT/US2000/004333
Other languages
French (fr)
Inventor
Tom L. Young
Michael G. Greene
Dennis R. Rice
Kelly L. Karlage
Sean P. Premeau
Original Assignee
Mbx Systems, Inc.
Priority date (The priority date is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the date listed.)
Filing date
Publication date
Priority claimed from AUPQ2706A external-priority patent/AUPQ270699A0/en
Priority claimed from AUPQ4144A external-priority patent/AUPQ414499A0/en
Application filed by Mbx Systems, Inc. filed Critical Mbx Systems, Inc.
Priority to CA 2363031 priority Critical patent/CA2363031C/en
Priority to BR0008962A priority patent/BR0008962A/en
Priority to APAP/P/2001/002267A priority patent/AP1870A/en
Priority to AU33705/00A priority patent/AU3370500A/en
Publication of WO2000048944A1 publication Critical patent/WO2000048944A1/en

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Classifications

    • CCHEMISTRY; METALLURGY
    • C01INORGANIC CHEMISTRY
    • C01GCOMPOUNDS CONTAINING METALS NOT COVERED BY SUBCLASSES C01D OR C01F
    • C01G23/00Compounds of titanium
    • C01G23/04Oxides; Hydroxides
    • C01G23/047Titanium dioxide
    • C01G23/053Producing by wet processes, e.g. hydrolysing titanium salts
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B11/00Obtaining noble metals
    • C22B11/06Chloridising
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B23/00Obtaining nickel or cobalt
    • C22B23/04Obtaining nickel or cobalt by wet processes
    • C22B23/0407Leaching processes
    • C22B23/0415Leaching processes with acids or salt solutions except ammonium salts solutions
    • C22B23/0423Halogenated acids or salts thereof
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B23/00Obtaining nickel or cobalt
    • C22B23/04Obtaining nickel or cobalt by wet processes
    • C22B23/0407Leaching processes
    • C22B23/0415Leaching processes with acids or salt solutions except ammonium salts solutions
    • C22B23/043Sulfurated acids or salts thereof
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B34/00Obtaining refractory metals
    • C22B34/10Obtaining titanium, zirconium or hafnium
    • C22B34/12Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08
    • C22B34/1236Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by wet processes, e.g. by leaching
    • C22B34/124Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by wet processes, e.g. by leaching using acidic solutions or liquors
    • C22B34/1245Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by wet processes, e.g. by leaching using acidic solutions or liquors containing a halogen ion as active agent
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B34/00Obtaining refractory metals
    • C22B34/10Obtaining titanium, zirconium or hafnium
    • C22B34/12Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08
    • C22B34/1236Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by wet processes, e.g. by leaching
    • C22B34/124Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by wet processes, e.g. by leaching using acidic solutions or liquors
    • C22B34/125Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by wet processes, e.g. by leaching using acidic solutions or liquors containing a sulfur ion as active agent
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B34/00Obtaining refractory metals
    • C22B34/10Obtaining titanium, zirconium or hafnium
    • C22B34/12Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08
    • C22B34/1236Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by wet processes, e.g. by leaching
    • C22B34/1259Obtaining titanium or titanium compounds from ores or scrap by metallurgical processing; preparation of titanium compounds from other titanium compounds see C01G23/00 - C01G23/08 obtaining titanium or titanium compounds from ores or scrap by wet processes, e.g. by leaching treatment or purification of titanium containing solutions or liquors or slurries
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B34/00Obtaining refractory metals
    • C22B34/30Obtaining chromium, molybdenum or tungsten
    • C22B34/34Obtaining molybdenum
    • CCHEMISTRY; METALLURGY
    • C22METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
    • C22BPRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
    • C22B7/00Working up raw materials other than ores, e.g. scrap, to produce non-ferrous metals and compounds thereof; Methods of a general interest or applied to the winning of more than two metals
    • C22B7/006Wet processes
    • YGENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
    • Y02TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
    • Y02PCLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
    • Y02P10/00Technologies related to metal processing
    • Y02P10/20Recycling

Definitions

  • Rutile (T ⁇ O 2 ) is a mineral used for many purposes. Amongst other uses, it is a source of titanium metal and a pamt pigment. Synthetic rutile is generally considered as any rutile created from another mineral, usually llmemte, that has at least 90% TiO,. High pu ⁇ ty rutile is 99.9%+ T ⁇ 0 2 . High pu ⁇ ty rutile generally carries a commercial value premium. llmemte (FeT ⁇ O 3 ) is most often converted to synthetic rutile by high temperature leaching with hydrochlo ⁇ c acid in an autoclave. Leachmg temperatures are generally between 800 to 900°C. Ferric chloride is sometimes used in these autoclave leaches to increase the reaction rates at the lower temperatures. Zoumei Jin et al (B. Mishra and G.J. Kiporous eds, In: Titanium Extraction and
  • the subject invention pertains to novel and highly efficient methods for leachmg valuable minerals, such as cobalt (Co), nickel (Ni), titanium (Ti), copper (Cu), molybdenum
  • One aspect of the present invention concerns methods for recove ⁇ ng titanium from ores.
  • One embodiment of the subject method uses an acidic solution, such as sulfuric acid, to leach titanium oxides from a mineral feed. Additional modifications and/or steps, including, for example, g ⁇ nding of the ore, addition of an alkali metal halide, addition of a carbon source, and adjustment of pressure and/or temperature, can be incorporated in the process.
  • a mineral feed is contacted with an acid and an alkali metal halide to leach titanium oxides from the feed.
  • High pu ⁇ ty titanium dioxide having a commercial premium over synthetic rutile can be produced using the methods of the subject invention.
  • the present invention provides a method for recovery of nickel and cobalt from a mineral feed by leachmg the feed with an acidic solution.
  • a mixture of sulfuric acid and an alkali metal halide are used to leach out cobalt and nickel from a late ⁇ te ore.
  • the subject methods can also be used to recover cobalt, nickel, copper, etc. by leachmg these elements from scrap metal.
  • the subject invention also concerns methods for recovering multiple metals or metal oxides in separate solutions.
  • ore is contacted with an acid solution, such as sulfuric acid.
  • Solid residue is then collected and contacted with an alkali metal halide and acid solution.
  • the subject method is used to recover copper separately from gold and silver. The copper is recovered p ⁇ ma ⁇ ly in the first acid solution, while the gold and silver are recovered in the alkali metal halide and acid solution.
  • Figure 2 shows the results of four consecutive one-hour leaches of titanium and iron from llmemte.
  • Figure 3 shows pulp density relationships m the leaching of titanium and iron from llmemte.
  • Figure 4 shows the results of expenments evaluating the effect of an alkali metal halide
  • Figure 5 shows the results of expenments evaluating the effects of grinding the ore on recovery rates.
  • Figure 6 shows the results of expenments evaluating the effect of adding a carbon source during the sulfunc acid leachmg process.
  • Figure 7 shows the results of expenments evaluating the effect of an alkali metal halide on the sulfu ⁇ c acid leachmg process of leachmg nickel from an initial late ⁇ te feed (Latente-1).
  • Figure 8 shows the results of expenments evaluating the effect of an alkali metal halide on the sulfu ⁇ c acid leachmg process of leach g cobalt from an initial late ⁇ te feed (Late ⁇ te-1).
  • Figure 9 shows the results of expenments evaluating the effect of an alkali metal halide on the sulfunc acid leach g process of leaching nickel from a second late ⁇ te feed (Latente-2).
  • Figure 10 shows the results of expenments evaluating the effects of an alkali metal halide on the sulfunc acid leachmg process of leachmg cobalt from a second late ⁇ te feed
  • the subject invention provides novel materials and methods useful for the recovery of minerals from ores.
  • An important component of the leachmg processes of the subject invention is the use of an acidic solution.
  • the acid is sulfu ⁇ c acid.
  • Sulfunc acid used in the leachmg procedures can be at a concentration ranging from about 20 grams per liter to about 500 grams per liter. In a preferred embodiment, the concentration of sulfunc acid ranges from about 150 grams per liter to about 250 grams per liter. Preferably, the concentration of sulfunc acid is approximately 200 grams per liter.
  • particularly preferred embodiments of the subject invention utilize additional factors including, for example, the use of an alkali metal halide, gnnding the ore, addition of a carbon source, and/or adjustment of the temperature at which the process is earned out.
  • the efficiency of the leachmg process can be improved by gnndmg the ore p ⁇ or to treatment.
  • the ore is ground so that it can pass through a 200 mesh sieve.
  • an alkali metal salt can be added to the leach solution to improve recovery.
  • the alkali metal salt can be for example, an alkali metal halide, alkali metal mt ⁇ te, alkali metal nitrate, alkali metal sulfite or alkali metal thiomte.
  • the metal halide can be, for example, NaCI, KCl, NaBr or KBr, or mixtures of one or more of these.
  • the metal sulfites can be, for example, sodium sulfite, sodium metabisulfite, sodium bisulfite, sodium dithionite, or other alkali metal or ammonium sulfite, metabisulfite, bisulfite or dithionite.
  • a further embodiment of the subject invention involves the use of a carbon source to improve recovery.
  • the carbon source can be, for example, graphite or activated carbon.
  • the source of this matenal can be, for example, from coconut shell or wood.
  • the present invention accordingly provides in one embodiment a method for recove ⁇ ng titanium ox ⁇ de(s) from a titanium and iron-containing mineral feed, the method including the steps of
  • the titanium ox ⁇ de(s) may be titanium dioxide.
  • the titamum-contammg mineral feed is typically post heavy mineral concentration products.
  • the feed will include titanium mineralization. Typical examples of this titanium mineralization are llmemte (FeT ⁇ O 3 ), leucoxene, perovskite (CaT ⁇ O 3 ) and titano magnetite.
  • the feed may m an alternative embodiment comp ⁇ se a bulk llmemte concentrate.
  • Other titanium-containing mineral feed matenals are contemplated within the scope of the invention.
  • the present invention provides in another separate embodiment a method for recove ⁇ ng synthetic rutile (T ⁇ O 2 ), from a mineral feed compnsmg llmemte (FeT ⁇ O 3 ), the method including the steps of:
  • the acidic solution preferably includes sulfunc acid.
  • the sulfunc acid used m the leachmg step is typically at a concentration m the range of from about 20 grams per liter to about 500 grams per liter.
  • the concentration of sulfunc acid is in the range of from about 150 grams per liter to about 250 grams per liter. Most preferably the concentration of sulfunc acid is about 200 grams per liter.
  • Other acids contemplated for use in step (a) of the present invention include, but are not limited to, a halide acid such as hydrochlonc acid or hydrobromic acid.
  • the typical concentration of halide acid used is in the range of from about 150 to about 350 grams per liter.
  • Step (a) is typically earned out in the presence of an alkali metal halide at a ratio of alkali metal halide to llmemte in the feed in the range of from about 1 : 1 to 2: 1.
  • the ratio of alkali metal halide is from about 1 : 1 to 1.5: 1. More preferably, the ratio is about 1.2: 1.
  • Suitable alkali metal ha des include, but are not limited to, NaCI, KCl or KBr or mixtures of one or more of these.
  • the alkali metal halide can be added directly to the leach solution.
  • the alkali metal halide can be combined with the feed p ⁇ or to introduction of the leachmg solution.
  • the feed may be subjected to a boildown pretreatment (i e., by boiling down to approximate dryness) m the presence of the alkali metal halide whereby the feed (e.g , llmemte surfaces) are coated with alkali metal halide p ⁇ or to leachmg.
  • a combination of the foregoing i.e , direct addition of alkali metal halide to the feed and combination of alkali metal halide with the feed p ⁇ or to leaching, can be used in the subject methods.
  • a proportion of the alkali metal halide is combined with the feed p ⁇ or to step (a) and a proportion of the alkali metal halide is added directly to the leach solution.
  • steps (a) and (b) may be conducted simultaneously or separately once solubihzation commences. It is particularly preferred to concurrently remove some of the pregnant solution from the leach residue to permit precipitation to take place away from the leach residue.
  • the precipitation step (b) can be regulated by adjustment of temperature and/or pH of the solution.
  • step (a) is earned out at a temperature in the range of from about 80°C to about 120°C and, preferably, is in the range of from about 90°C to about 110°C.
  • the operating temperature for step (a) is about 100°C.
  • the leach solution m step (a) has a solids content of up to about 60% by weight.
  • the leach solution has a solids content of from about 10% to about 40%.
  • the feed may be ground mto finer particles.
  • the feed may be subjected to fine g ⁇ ndmg.
  • the majority of particles in the feed are capable of passing through a 75 micron sieve after g ⁇ ndmg.
  • a source of carbon may be provided m the subject method.
  • the carbon may be in the form of any commercially available carbon source including, for example, activated carbon, coal, coke, charcoal or graphite.
  • a preferred source of carbon is activated carbon derived from coconut shell.
  • the ratio of carbon to feed (e g., llmemte) is typically between about 0.01:1 to 1 : 1.
  • Methods according to the present invention may be earned out at or above atmospheric pressure.
  • the typical elevated pressures and temperatures at which the present methods may be performed are in the range of from about 1 bar to about 30 bar.
  • pressures are in the range of from about 1 bar to about 5 bar.
  • Temperatures used in the subject methods range from about 100°C to about 235°C.
  • temperatures range from about 100°C to about 150°C.
  • the leach residue produced from step (a) can be subjected to further leaching to solubilize undissolved iron and or titanium in the residue.
  • the further leachmg can be performed using fresh acidic solution.
  • spent leach liquor or a combination of fresh acidic solution and spent leach liquor can be used.
  • step (a) of the subject method can be performed in the presence of ferrous and/or feme ions to promote dissolution of the iron mineralization.
  • Ferrous ions will generally be present in recirculated process plant solutions
  • iron may be removed from the leachant solution using standard techniques, such as precipitation.
  • the purpose is to remove soluble iron from any process solutions.
  • Solvent extraction, ion exchange, reverse osmosis or other techniques can also be used to remove soluble iron.
  • the leach time for this embodiment is generally relatively long, and typically is in the range of from about 50 to about 120 hours. Preferably, leach time is from about 60 to about 100 hours.
  • the operating conditions are much milder than conventional autoclave techniques, leading to large capital and operating cost advantages. Sulfu ⁇ c acid and alkali metal halides are easier to handle than the hydrochloric acid used m the Zoumei Jm et al. process refe ⁇ ed to above.
  • the present invention provides m another separate embodiment, a method for recovenng titanium from a titanium and iron-containing mineral feed, the method including the steps of (a) solubihzmg titanium and iron by leachmg the feed with an acidic solution in the presence of an alkali metal halide and a source of activated carbon;
  • the present invention provides in another separate embodiment a method for recove ⁇ ng titanium from a mineral feed comp ⁇ smg llmemte (FeT ⁇ O 3 ), the method including the steps of
  • the present invention provides m another separate embodiment a method for recovenng titanium ox ⁇ de(s) from a mineral feed comprising llmemte (FeT ⁇ O 3 ), the method including the steps of : (a) leachmg the llmemte with an acidic solution at a temperature m the range of from about 80 to 120°C m the presence of an alkali metal halide for a predetermined time, the leach solution containing up to about 60% by weight solids to produce a leachant solution containing iron and titanium ions; (b) separating the iron from the titanium in the leachant solution; and (c) recovering the separated titanium as T ⁇ 0 2 .
  • reaction time of step (a) of this embodiment is up to about an hour.
  • reaction time of step (a) is up to about half an hour. More preferably, the reaction time is m the range of from about 5 to about
  • step (a) above may be repeated to solubilize unleached titanium in the residue obtained following step (a) in order to obtain cumulative maximum solubility of titanium.
  • Fresh acidic solution and alkali metal halide can be used when step (a) is repeated.
  • Step (a) may m one embodiment comprise a type of countercurrent leach circuit.
  • the acidic solution in this embodiment can be supplemented with hydrochlonc acid in one or more steps of a repeated leach sequence to assist m enhancing the titanium solubility profile.
  • the present invention provides a method for recove ⁇ ng titanium from a titanium and iron-containmg mineral feed, the method including the steps of:
  • the halide acid used in step (a) can be, for example, hydrochloric acid or hydrobromic acid.
  • concentration of halide acid used can be in the range of from about 150 to about 350 grams per liter acid.
  • Any precipitated titanium reporting to the leach residue of this embodiment may be recovered m subsequent leachmg operations.
  • the present invention provides m another separate embodiment a method for recove ⁇ ng titanium from a feed comp ⁇ smg finely ground llmemte (FeT ⁇ 0 3 ), the method including the steps of
  • step (b) repeating step (a); (c) separating at least some of the pregnant solution from the leach residue;
  • the present invention provides multistage leaching of iron and titanium from an iron and titamum-bearmg mineral feed, the method comp ⁇ sing the following steps. (a) contacting the feed mate ⁇ al with an acid — alkali halide solution for a penod of time sufficient to solubilize the titanium but not so long as to allow the titanium to appreciably precipitate;
  • step (b) separating the pulp from the leach liquor; (c) contacting the pulp with fresh leach liquor and repeating steps (a) and (b) until all economically feasible titanium is leached; and (d) selectively recove ⁇ ng the titanium and iron m separate stages from the leach solutions by precipitation, solvent extraction or other means.
  • the conditions of step (a) can involve percent solids on a weight/weight basis of between about 1 percent and about 60 percent.
  • the typical percent solids are in the range of from about 10% to 40%
  • the solids may be ground to fine size to facilitate leachmg, typically so that the feed passes a 73 micron sieve
  • the acid used is most typically sulfunc acid.
  • the acid concentration can range from about 20 to about 300 grams per liter acid. Most typically the acid concentration ranges from about 150 to 230 gram per liter.
  • the alkali halide can be any alkali halide.
  • the alkali halide is NaCI, KCl,
  • the concentration of alkali halide can range from about 50 grams per liter to about 400 grams per liter. Preferably, the alkali halide concentration is about 100 to about 200 grams per liter.
  • the leaching is most typically carried out at about room pressure
  • the temperature can be between about 40°C and about 110°C at room pressure
  • leachmg temperature is between about 90 °C and about 105°C.
  • Leaching at room pressure will typically be performed m a leach vessel with a condenser to limit the loss of halide acid generated m the leach solution
  • the titanium is allowed to reach a concentration as high as possible before it begins to re- precipitate onto the leach feed matenal. This is typically slightly over four (4) grams of titanium per liter of solution.
  • the leach time to accomplish this solubihzation will depend on the va ⁇ ous aforementioned parameters but will usually range from about 10 minutes to 1 hour
  • the method of solid — liquid separation in step (b) can be any method that makes a good separation of the solids from the leach liquor in a relatively short time. These include methods such as cyclones, filters, centnfuges, magnetic separators, and settlers. The list is not meant to exclude any unnamed method.
  • the fresh leach liquor in step (c) can be leach liquor from which the titanium content has been reduced or eliminated.
  • the iron content of liquor should be controlled so that no precipitation of an iron compound occurs dunng the leachmg
  • the titanium can be totally or partially removed from the leach liquor in step (d) by the method that makes the most economic sense for any given plant.
  • the methodology available includes, but is not limited to, precipitation by seeding or pH adjustment, crystallization, solvent extraction, and ion exchange.
  • the iron can be removed in a similar fashion in a step before or after the titanium recovery. Titanium and iron are recovered as separate products, in separate stages. The titanium would be recovered as a titanium salt, most typically T ⁇ O 2 . The iron would most typically be recovered as an iron salt such as ferrous chlo ⁇ de or sulphate.
  • the present invention also concerns methods for the recovery of other minerals, such as nickel, cobalt, copper, molybdenum, lead, zmc, gold or silver from ore, soil, concentrate, slag or residue.
  • other minerals such as nickel, cobalt, copper, molybdenum, lead, zmc, gold or silver from ore, soil, concentrate, slag or residue.
  • a method is provided for the dissolution of nickel and cobalt from a nickel, cobalt and iron-containing mineral feed, the method compnsmg solubihzmg the nickel, cobalt and iron m the feed by leachmg the feed with an acidic solution.
  • an alkali metal salt can be added to the leach solution to improve recovery.
  • the alkali metal salt can be for example, an alkali metal halide, alkali metal nit ⁇ te, alkali metal nitrate, alkali metal sulfite or alkali metal thiomte.
  • the metal halide can be, for example, NaCI, KCl, NaBr or KBr, or mixtures of one or more of these.
  • the metal sulfites can be, for example, sodium sulfite, sodium metabisulfite, sodium bisulfite, sodium dithionite, or other alkali metal or ammonium sulfite, metabisulfite, bisulfite or dithionite.
  • the method of the invention can be conducted at above ambient temperatures and at or above atmospheric pressures prior to metal extraction by precipitation, solvent extraction or other means
  • the nickel and cobalt-contammg mineral feed is typically post beneficiation by comminution and thickening products.
  • a typical example of nickel and cobalt mineralization is a latente ore.
  • the feed may compnse a bulk latente concentrate.
  • One embodiment of the present method provides for recovering nickel and cobalt from a mineral feed comprising late ⁇ te, the method including the step of solubihzmg nickel and cobalt and iron in the late ⁇ te by leachmg the latente with an acidic solution in the presence of an alkali metal halide at a temperature not exceeding about 150°C at normal pressures prior to nickel and cobalt extraction by established precipitation, solvent extraction or other means.
  • the acidic solution contains sulfuric acid.
  • the sulfuric acid used in the leachmg step is typically at a concentration in the range of from about 20 grams per liter to about 500 grams per liter.
  • the concentration of sulfu ⁇ c acid is in the range of from about 150 grams per liter to about 250 grams per liter.
  • the concentration of sulfunc acid is about 200 grams per liter.
  • Other acids contemplated for use in the present invention include halide acids, for example, hydrochlonc acid or hydrobromic acid.
  • the typical concentration of halide acid used is in the range from about 50 to about 350 grams per liter acid.
  • the process is typically earned out m the presence of an alkali metal halide at a ratio of alkali metal halide to late ⁇ te in the feed m the range of from about 0.05:1 to about 4:1.
  • the ratio is about 0J : 1 , and most preferably about 0.2: 1.
  • the alkali metal salt may be added directly to the leach solution.
  • the alkali metal salt is combined with the feed p ⁇ or to introduction of the leachmg solution.
  • the feed may be subjected to a boildown pre-treatment (i.e., by boiling down to approximate dryness) in the presence of the alkali metal salt whereby the feed (e g , latente) surfaces are coated with alkali metal salt p ⁇ or to leachmg.
  • the solution of alkali salt may be sprayed on a heap of latentic ore and allowed to evaporate.
  • a proportion of the alkali metal salt is combined with the feed p ⁇ or to solubihzation and a proportion of the alkali metal salt is added directly to the leach solution. It is particularly preferred to concurrently remove some of the pregnant solution from the leach residue to permit separation of the nickel and cobalt to take place away from the leach residue.
  • the process is carried out at a temperature in the range of from about 80 °C to about 120°C.
  • the temperature is m the range of from about 90 °C to about 110°C
  • a typical operating temperature for the process is about 100°C.
  • the leach solution in the subject process preferably has a solids content of up to about 60% by weight.
  • the leach solution has a solids content of from about 10 to 40%.
  • the feed can be ground into smaller particles. It is preferred that the feed be subjected to fine gnndmg.
  • the majority of particles in the feed are capable of passing through 75 micron sieve. Typically, at least 75% of the particles m the feed are of a size that can pass through 75 micron sieve openings.
  • Methods according to the present invention may be earned out at or above atmosphenc pressure.
  • the typical elevated pressures and temperatures at which methods according to the invention may be performed are in the range of from about 1 bar to about 30 bar.
  • pressures are m the range of from about 1 bar to about 5 bar and temperatures range from about 100°C to about 235 °C.
  • m the range of from about 100°C to about 150°C.
  • the leach residue produced by the present process may be subjected to further leachmg to solubilize undissolved iron and/or nickel and cobalt m the residue.
  • the further leachmg can be performed using fresh acidic solution.
  • spent leach liquor, or a combination of fresh acidic solution and spent leach liquor may be used in the process.
  • the process may be performed m the presence of ferrous and/or feme ions to promote dissolution of the iron mineralization.
  • Ferrous ions will generally be present in recirculated process plant solutions.
  • a typical reaction time for the process of this embodiment is up to about six hours.
  • the reaction time is up to about two hours. More preferably, the reaction time is in the range of from about 15 minutes to about 3 hours. It has been observed that nickel and cobalt solubility reaches a peak dunng reaction times of approximately that length.
  • a person of ordinary skill in the art can vary leach time so as to leach less of an unwanted species such as manganese or iron at the expense of some cobalt and nickel recovery.
  • the process above may be repeated to solubilize unleached nickel or cobalt m the residue m order to obtain cumulative maximum solubility of nickel and cobalt.
  • Fresh acidic solution and alkali metal halide may be used when the process is repeated.
  • the process may in one embodiment comprise a type of countercurrent leach circuit.
  • the acidic solution may in this embodiment be supplemented with hydrochloric acid in one or more steps of a repeated leach sequence to assist in enhancing the nickel or cobalt solubility profile.
  • a metal halide salt may be used either to precondition an aqueous slurry or it may be sprayed onto the feed material and allowed to evaporate prior to contacting with sulfunc acid.
  • This flash leachmg process utilizes the selective nature of the leach to achieve a cobalt nch solution containing only minor amounts of nickel, manganese, iron, etc.
  • super alloy scrap and other recycled metal alloys may be leached by treating with a halide salt of the metal and sulfunc acid.
  • concentrations of the metal halide salt and the sulfunc acid will be dependent upon the specific scrap mixture.
  • This embodiment can be utilized to selectively leach specific metals or to place all the metals into solution.
  • This embodiment may also be used to solubilize radio-nucleosides of such metal as nickel from a radiated scrap. Oxygen or other oxidizing gasses such as chlo ⁇ ne can be added to the system to oxidize the metal.
  • the alkali metal halide may be substituted with a sulfur-based reducing chemical.
  • a sulfur-based reducing chemical For example, sodium sulfite, sodium metabisulfite, sodium bisulfite, sodium dithionite, or other alkali metal or ammonium sulfite, metabisulfite. bisulfite or dithionite can be used in place of the alkali metal halide.
  • sulfur based reducing chemicals will facilitate the reduction of the transition metal, opening the ore up to attack by the sulfu ⁇ c acid. The metal of economic interest need not always be the one reduced.
  • Alkali metal nitrates or nitntes may be used with sulfunc acid to leach most metals. These techniques may be used to leach metals from sulfide minerals or from scrap, residue, slags, concentrates, or soils.
  • the process utilizing a metal halide salt and sulfunc acid may be used, with minor modifications, in currently existing counter current decantation (CCD) circuits. Such an embodiment would utilize fresh feed matenal to achieve neutralization to a pH adequate to retain iron in solution.
  • CCD counter current decantation
  • the resultant leach liquor may be further neutralized to precipitate iron as a hydroxide m the presence of a binding material. The iron precipitate may then be partially dned and pelletised to produce pig iron feed stocks.
  • the method of solid-liquid separation can be any method that produces a good separation of the solids from the leach liquor in a relatively short time. These include, but are not limited to, methods such as cyclones, filters, centrifuges, magnetic separators, and settlers.
  • the nickel or cobalt can be totally or partially removed from the leach liquor by the method that makes the most economic sense for any given plant.
  • the methodology available includes, but is not limited to, precipitation by seeding or pH adjustment, crystallization, solvent extraction, and ion exchange.
  • the subject invention also concerns methods for recovering multiple metal or metal oxides in separate solutions.
  • Mineral species of economic value are often associated with species that consume the chemical reagents that are used to leach them. Sometimes even though the consuming species is of economic value, the overall leach becomes uneconomic. The most common example of this is copper-gold ores. Following are examples which illustrate procedures for practicing the invention. These examples should not be construed as limiting. All percentages are by weight and all solvent mixture proportions are by volume unless otherwise noted.
  • the Ti appears to be leached withm one hour.
  • a 100 gram quantity of llmemte with a head assay of 34.0%> Fe and 27.0% Ti, and particle size such that 100% of the particles pass through a 75 micron screen was leached for 72 hours at 100° C m 1 liter of 200 gram per liter H 2 SO 4 — 120 gram per liter alkali metal halide solution.
  • a 100 gram quantity of activated carbon was also present in the leach solution.
  • the leach liquor was monitored periodically for Ti and Fe content. The results of the experiment are shown in Table 1. Titanium is dissolved then observed to subsequently precipitate.
  • Example 2 Using the data in Example 1 allowed the development of a new leach procedure for llmemte.
  • the procedure comp ⁇ ses leachmg llmemte for one hour or less and then contacting it with fresh leach solution. In this manner both the iron and titanium is leached together. This was tested using the same conditions as m the 96-hour test. The results of four (4) consecutive one-hour leaches on the same ore sample are shown m Figure 2. As can be seen, approximately the same amount of iron and llmemte was leached in each step.
  • the ordinarily skilled artisan having the benefit of the teachings desc ⁇ bed herein, can determine the proper reagent concentration, temperature, particle size of the ore, whether to include carbon and its form (e.g., activated carbon or graphite), or atmosphenc pressure (typically ⁇ 3 atmospheres) that is optimum for a particular ore.
  • the technique of separating the Ti as T ⁇ 0 2 with short leach times followed by precipitation of T ⁇ O 2 is also applicable to other leach systems such as the hydrochlonc acid leach system.
  • Experiment A comprises a leach solution of 60 grams alkali metal halide, 100 grams H 2 ,SO 4 , and 350 grams of H 2 O heated to 100 C C m Erlenmeyer flasks on a stir ⁇ ng hotplate to which is added 50 grams of minus 75 microns particle size llmemte resulting m a 9% pulp density.
  • Experiment B comprises a leach solution of 60 grams alkali metal halide, 100 grams H 2 S0 4 , and 350 grams of H 2 0 heated to 100°C in Erlenmeyer flasks on a stir ⁇ ng hotplate to which is added 100 grams of minus 75 microns particle size ilmemte resulting m a 16% pulp density.
  • the ilmemte had an assay head of 30% titanium and 34% iron.
  • the following procedure steps are applied separately to Expe ⁇ ment A and Expe ⁇ ment B:
  • Step 1 A condenser is placed on the Erlenmeyer containing the slurry compnsmg the presc ⁇ bed solution and ilmemte feed; Step 2. The slurry is stirred vigorously with a magnetic stirrer for 30 minutes with the temperature maintained at 100°C; Step 3. The Erlenmeyer and contents are cooled for a couple of minutes in a room temperature water bath; Step 4. The Erlenmeyer solution is decanted into a cent ⁇ fuge rube and centrifuged at 4,000 rpm for 5 minutes; Step 5. The liquor in the ccntnfugc tube is decanted and separated from the solids into a sample bottle, volume and weight determined and retained for further test work including analysis; Step 6.
  • Step 8 The procedure is continued by repeating Steps 1 through 7 inclusive, a total of seven times and thus equating to a total leach duration of 4 hours; Step 9.
  • the post cent ⁇ fugmg liquors collected at each repetition of Step 5 are individually subsampled and analysed for titanium and iron;
  • Step 10 Calculations are conducted to determine titanium and iron contents of both solids and liquors and comparisons made with the respective elemental assay values of the ilmemte ore feed;
  • Step 11 The individual liquors remaining after the subsamplmg conducted in Step 9 are combined in a flask and subsampled and analysed for titanium and iron; Step 12.
  • the titanium can be totally or partially removed from the leach liquor by the method that makes the most economic sense for any given plant.
  • the methodologies available include, but are not limited to, precipitation by seeding or pH adjustment, crystallization, solvent extraction, and ion exchange.
  • Example 3 Effect of Alkali Metal Halide
  • the salt which was used was NaCI at 0%, 5%, 15% and 25%
  • Example 4 Effect of Gnndmg of Ore
  • grinding of the ore can be used to increase the reaction rate of leaching iron from ilmenite. This is shown in Figure 5 and Table 4. Both tests were performed using a 100 gram quantity of ilmenite placed in one liter of 200 grams per liter sulfuric acid and 150 grams per liter alkali metal halide solution heated to 100°C. The experiments were conducted on two samples of the same ilmenite feed. One experiment used course ilmenite (100% retained on a 75 micron screen) and the other experiment used fine ilmenite (100% passing through a 75 micron screen). The slurry was vigorously stirred for 72 hours and the iron concentration periodically monitored. The ground ore (finer particle sized samples) had faster early and late leach kinetics than the unground ore (coarser particle sized sample). The kinetics of the ore during the 5 to 25 hour time period was similar in both cases.
  • Example 6 Leachmg of Copper and Nickel from Latente Ore with a Sulfuric Acid-Hahde- Carbon System
  • This ore has an assay head of 2.36% Co, 1.26% Ni, 11.00% Fe, 10.80% Mn.
  • a sample of 100 grams of ground, -200 mesh ore was first treated with 200 grams of NaCI dissolved in 650 grams of water The water was evaporated on a hot plate. This procedure is a speeded up simulation of spaymg a heap of ore with a salt solution and letting it evaporate naturally. The ore-salt solids were then slur ⁇ ed in 200 grams of sulfu ⁇ c acid in 700 grams of water solution. The stirred slurry was brought to 100° C on a stirnng hot plate, and then 100 grams of +65 mesh, coconut shell, activated carbon was added.
  • Example 7 Effect of Alkali Metal Halide on the Leachmg of Nickel and Cobalt from Latente 1 Expenments were conducted on two samples of 100 grams of late ⁇ te- 1 feed, compnsing
  • Step 1 A condenser is placed on the Erlenmeyer containing the slurry comp ⁇ sing the prescnbed solution and late ⁇ te feed;
  • Step 2 The slurry is stirred vigorously with a magnetic stirrer for the duration of the test with the temperature maintained at 100°C; Step 3.
  • the test is sampled at predetermined times, eg., 15 minutes, 30 minutes, etc., by pipetting 10 ml of the hot slurry from the Erlenmeyer into a centnfuge tube and centrifuge at 4,000 ⁇ m for 5 minutes;
  • Step 4 The centnfuged timed leach solution is transferred mto a sample tube for later analysis; Step 5. 10 ml of make-up leach solution is used to wash the centnfuged residue back into the Erlenmeyer, while the Erlenmeyer continues to be agitated at 100°C on the hot plate; Step 6. At the end of the test (e.g. , 6 hours) the contents of the Erlenmeyer is poured into two centnfuge tubes, using an additional very small amount of distilled water to wash out any residue remaining on the mside lip of the Erlenmeyer, and then centrifuged; Step 7.
  • centnfuged liquid contents from both centnfuge tubes is decanted into a graduated cylinder and allow to cool; Step 8. Then having read the volume of PLS solution, approximately 20 ml is transferred into a sample tube and analysed for nickel and cobalt; Step 9. Calculations are conducted to determine nickel and cobalt contents of the liquors and compa ⁇ sons made with the respective elemental assay values of the late ⁇ te ore feed; Step 10. Nickel and cobalt can be totally or partially removed from the leach liquor by the method that makes the most economic sense for any given plant.
  • the methodologies available va ⁇ ously include, but are not limited to, precipitation of metallic salts by seeding, pH adjustment, or crystallisation; solvent extraction and electrowmnmg of elemental metal; and ion exchange.
  • This example shows the leachmg of silver from a copper refinery pilot plant's slimes
  • the test was conducted at 100°C with 200 gram per liter sulfu ⁇ c acid and 200 gram per liter NaCI. Samples of 50 grams of slimes were leached in 500 miUiliters of solution. The leachmg was conducted for 48 hours. The results are shown m the Table 9.
  • Example 11 Two Stage Leaching of Different Metals into Two Separate Leach Liquors An oxide copper ore sample, ground to minus 200 mesh, with a head grade of 0.91%>
  • a person skilled in the art, having the benefit of the teachings of this disclosure, can adjust the acid concentration and temperature to achieve complete recovery of the copper in the first stage while maintaining excellent recovery of the Au and Ag in the second stage.
  • the subject method can also be readily adapted to heap leaching.

Abstract

Disclosed and claimed are efficient methods for leaching minerals from ores using an acidic solution such as sulfuric acid. Additional factors which can improve mineral recovery include the use of an alkali metal halide, grinding the ore, addition of a carbon source, and/or adjustment of the temperature at which the process is carried out. Minerals such as titanium, iron, nickel, cobalt, silver and gold may be recovered by the methods of the present invention.

Description

DESCRIPTION
IMPROVED METHODS FOR LEACHING OF ORES
Cross-Reference to Related Application
This application claims the benefit of the filing date of provisional application USSN 60/120,820 filed February 19, 1999.
Background of the Invention Oxides of cobalt (Co), nickel (Ni), titanium (Ti), copper (Cu), molybdenum (Mo), lead
(Pb), zmc (Zn), gold (Au), and silver (Ag) are important minerals Various methods exist for recovering these compounds from the ores where they are found. For example, autoclave methods are often used to recover Co, Ni and Ti oxides. These methods are capital and labor intensive. Mo oxide has been leached by hydrochloπc acid methods. Cyanide, thiosulfate, thiourea and ha des are used in leaching Au and Ag metals and oxides. Cu, Zn and Pb can be leached with sulfuπc acid.
Rutile (TιO2) is a mineral used for many purposes. Amongst other uses, it is a source of titanium metal and a pamt pigment. Synthetic rutile is generally considered as any rutile created from another mineral, usually llmemte, that has at least 90% TiO,. High puπty rutile is 99.9%+ Tι02. High puπty rutile generally carries a commercial value premium. llmemte (FeTιO3) is most often converted to synthetic rutile by high temperature leaching with hydrochloπc acid in an autoclave. Leachmg temperatures are generally between 800 to 900°C. Ferric chloride is sometimes used in these autoclave leaches to increase the reaction rates at the lower temperatures. Zoumei Jin et al (B. Mishra and G.J. Kiporous eds, In: Titanium Extraction and
Processing, The Mineral, Metals & Materials Society (1997) pg 122-128) reported that 4 to 6 Normal hydrochloπc acid at 110 to 140°C, will dissolve the iron (Fe) from llmemte from the Sichuan province of China m 6 hours. They found that the reaction rate is 0.4 order with respect to initial Fe+2 concentration. They postulate a surface reaction control model with an apparent activation energy of 56.97 kiloj oules per mol .
Conventional autoclave technology is capital, maintenance and energy intensive. The process disclosed in Zoumei Jin et al. process involves the use of large amounts of hydrochloπc acid, which is expensive, difficult to handle and requires special stainless steel equipment. There is a clear need for more efficient processes for leachmg of ores to obtain valuable minerals. Cyanide is the most commonly used leachant for gold. Two molecules of cyanide complex with every molecule of gold. Copper also complexes with cyanide, but it takes 4 molecules of cyanide for every copper molecule. Copper is often present in copper gold ores in the one tenth to one percent range. Gold m these ores is in the one to 10 parts per million range. The copper consumes so much cyanide it needs to be recovered by hydrogen cyanide distillation, an expensive and dangerous operation. Systems have been proposed where sulfuπc acid is used to leach the copper first. Then, the heap or batch is neutralized. Cyanide can then be used to leach the gold. The problem, of course, is the expense of neutralization. In heap operations, the additional worry of incomplete neutralization is present. In other ores, the gangue, or unwanted mateπal, can be an acid consumer. Copper oxide m limestone rich rock is an example.
Brief Summary of the Invention The subject invention pertains to novel and highly efficient methods for leachmg valuable minerals, such as cobalt (Co), nickel (Ni), titanium (Ti), copper (Cu), molybdenum
(Mo), lead (Pb), zinc (Zn), gold (Au), and silver (Ag) from ores.
One aspect of the present invention concerns methods for recoveπng titanium from ores. One embodiment of the subject method uses an acidic solution, such as sulfuric acid, to leach titanium oxides from a mineral feed. Additional modifications and/or steps, including, for example, gπnding of the ore, addition of an alkali metal halide, addition of a carbon source, and adjustment of pressure and/or temperature, can be incorporated in the process. In a preferred embodiment, a mineral feed is contacted with an acid and an alkali metal halide to leach titanium oxides from the feed. High puπty titanium dioxide having a commercial premium over synthetic rutile can be produced using the methods of the subject invention. Another aspect of the present invention concerns methods for recovering transition metals other than titanium from ores In one embodiment, the present invention provides a method for recovery of nickel and cobalt from a mineral feed by leachmg the feed with an acidic solution. In an exemplified embodiment, a mixture of sulfuric acid and an alkali metal halide are used to leach out cobalt and nickel from a lateπte ore. The subject methods can also be used to recover cobalt, nickel, copper, etc. by leachmg these elements from scrap metal.
The subject invention also concerns methods for recovering multiple metals or metal oxides in separate solutions. In one embodiment, ore is contacted with an acid solution, such as sulfuric acid. Solid residue is then collected and contacted with an alkali metal halide and acid solution. In an exemplified embodiment, the subject method is used to recover copper separately from gold and silver. The copper is recovered pπmaπly in the first acid solution, while the gold and silver are recovered in the alkali metal halide and acid solution.
Bnef Descnption of the Drawings Figure 1 shows the kinetics of leachmg titanium and iron from llmemte.
Figure 2 shows the results of four consecutive one-hour leaches of titanium and iron from llmemte.
Figure 3 shows pulp density relationships m the leaching of titanium and iron from llmemte. Figure 4 shows the results of expenments evaluating the effect of an alkali metal halide
(NaCI) on the sulfuπc acid leachmg process
Figure 5 shows the results of expenments evaluating the effects of grinding the ore on recovery rates.
Figure 6 shows the results of expenments evaluating the effect of adding a carbon source during the sulfunc acid leachmg process.
Figure 7 shows the results of expenments evaluating the effect of an alkali metal halide on the sulfuπc acid leachmg process of leachmg nickel from an initial lateπte feed (Latente-1).
Figure 8 shows the results of expenments evaluating the effect of an alkali metal halide on the sulfuπc acid leachmg process of leach g cobalt from an initial lateπte feed (Lateπte-1). Figure 9 shows the results of expenments evaluating the effect of an alkali metal halide on the sulfunc acid leach g process of leaching nickel from a second lateπte feed (Latente-2).
Figure 10 shows the results of expenments evaluating the effects of an alkali metal halide on the sulfunc acid leachmg process of leachmg cobalt from a second lateπte feed
(Lateπte-2).
Detailed Disclosure of the Invention The subject invention provides novel materials and methods useful for the recovery of minerals from ores. An important component of the leachmg processes of the subject invention is the use of an acidic solution. In one embodiment, the acid is sulfuπc acid. Sulfunc acid used in the leachmg procedures can be at a concentration ranging from about 20 grams per liter to about 500 grams per liter. In a preferred embodiment, the concentration of sulfunc acid ranges from about 150 grams per liter to about 250 grams per liter. Preferably, the concentration of sulfunc acid is approximately 200 grams per liter.
In addition to using sulfunc acid solutions in the leachmg processes of the subject invention, particularly preferred embodiments of the subject invention utilize additional factors including, for example, the use of an alkali metal halide, gnnding the ore, addition of a carbon source, and/or adjustment of the temperature at which the process is earned out.
In accordance with the subject invention, the efficiency of the leachmg process can be improved by gnndmg the ore pπor to treatment. In a prefened embodiment, the ore is ground so that it can pass through a 200 mesh sieve.
In a further embodiment, an alkali metal salt can be added to the leach solution to improve recovery. The alkali metal salt can be for example, an alkali metal halide, alkali metal mtπte, alkali metal nitrate, alkali metal sulfite or alkali metal thiomte. The metal halide can be, for example, NaCI, KCl, NaBr or KBr, or mixtures of one or more of these. The metal sulfites can be, for example, sodium sulfite, sodium metabisulfite, sodium bisulfite, sodium dithionite, or other alkali metal or ammonium sulfite, metabisulfite, bisulfite or dithionite. The ordinarily skilled artisan, having the benefit of the teachings disclosed herein, can readily determine those alkali metal salts that can be used in conjunction with the particular acid solution used in the solubihzation step of the process. A further embodiment of the subject invention involves the use of a carbon source to improve recovery. The carbon source can be, for example, graphite or activated carbon. The source of this matenal can be, for example, from coconut shell or wood.
The present invention accordingly provides in one embodiment a method for recoveπng titanium oxιde(s) from a titanium and iron-containing mineral feed, the method including the steps of
(a) solubihzmg titanium and iron by leachmg the feed with an acidic solution m the presence of an alkali metal halide;
(b) selectively precipitating titanium oxιde(s), and
(c) recovenng titanium oxιde(s). Typically, the titanium oxιde(s) may be titanium dioxide.
The titamum-contammg mineral feed is typically post heavy mineral concentration products. The feed will include titanium mineralization. Typical examples of this titanium mineralization are llmemte (FeTιO3), leucoxene, perovskite (CaTιO3 ) and titano magnetite. The feed may m an alternative embodiment compπse a bulk llmemte concentrate. Other titanium-containing mineral feed matenals are contemplated within the scope of the invention.
The present invention provides in another separate embodiment a method for recoveπng synthetic rutile (TιO2), from a mineral feed compnsmg llmemte (FeTιO3), the method including the steps of:
(a) solubihzmg titanium and iron by leachmg the llmemte with an acidic solution in the presence of an alkali metal halide, (b) selectively precipitating titanium oxιde(s), and
(c) recovenng titanium oxide as TιO2.
In step (a) of the method, the acidic solution preferably includes sulfunc acid. The sulfunc acid used m the leachmg step is typically at a concentration m the range of from about 20 grams per liter to about 500 grams per liter. In a preferred embodiment, the concentration of sulfunc acid is in the range of from about 150 grams per liter to about 250 grams per liter. Most preferably the concentration of sulfunc acid is about 200 grams per liter. Other acids contemplated for use in step (a) of the present invention include, but are not limited to, a halide acid such as hydrochlonc acid or hydrobromic acid. The typical concentration of halide acid used is in the range of from about 150 to about 350 grams per liter.
Step (a) is typically earned out in the presence of an alkali metal halide at a ratio of alkali metal halide to llmemte in the feed in the range of from about 1 : 1 to 2: 1. Preferably, the ratio of alkali metal halide is from about 1 : 1 to 1.5: 1. More preferably, the ratio is about 1.2: 1. Suitable alkali metal ha des include, but are not limited to, NaCI, KCl or KBr or mixtures of one or more of these.
In the methods of the present invention, the alkali metal halide can be added directly to the leach solution. Alternatively, the alkali metal halide can be combined with the feed pπor to introduction of the leachmg solution. In this case, the feed may be subjected to a boildown pretreatment (i e., by boiling down to approximate dryness) m the presence of the alkali metal halide whereby the feed (e.g , llmemte surfaces) are coated with alkali metal halide pπor to leachmg. Optionally, a combination of the foregoing, i.e , direct addition of alkali metal halide to the feed and combination of alkali metal halide with the feed pπor to leaching, can be used in the subject methods. Thus, for example, a proportion of the alkali metal halide is combined with the feed pπor to step (a) and a proportion of the alkali metal halide is added directly to the leach solution. Typically, steps (a) and (b) may be conducted simultaneously or separately once solubihzation commences. It is particularly preferred to concurrently remove some of the pregnant solution from the leach residue to permit precipitation to take place away from the leach residue. In this way, the precipitate may be restncted from coating the leach residue which could potentially decrease the efficiency of the process. In one embodiment, the precipitation step (b) can be regulated by adjustment of temperature and/or pH of the solution. Typically, step (a) is earned out at a temperature in the range of from about 80°C to about 120°C and, preferably, is in the range of from about 90°C to about 110°C. In a preferred embodiment, the operating temperature for step (a) is about 100°C. In one embodiment, the leach solution m step (a) has a solids content of up to about 60% by weight. Preferably, the leach solution has a solids content of from about 10% to about 40%.
To facilitate more rapid leachmg, the feed may be ground mto finer particles. In a preferred embodiment, the feed may be subjected to fine gπndmg. Preferably, the majority of particles in the feed are capable of passing through a 75 micron sieve after gπndmg.
Optionally, a source of carbon may be provided m the subject method. The carbon may be in the form of any commercially available carbon source including, for example, activated carbon, coal, coke, charcoal or graphite. A preferred source of carbon is activated carbon derived from coconut shell. The ratio of carbon to feed (e g., llmemte) is typically between about 0.01:1 to 1 : 1.
Methods according to the present invention may be earned out at or above atmospheric pressure. When elevated pressures are used, the typical elevated pressures and temperatures at which the present methods may be performed are in the range of from about 1 bar to about 30 bar. Preferably, pressures are in the range of from about 1 bar to about 5 bar. Temperatures used in the subject methods range from about 100°C to about 235°C. Preferably temperatures range from about 100°C to about 150°C.
The leach residue produced from step (a) can be subjected to further leaching to solubilize undissolved iron and or titanium in the residue. The further leachmg can be performed using fresh acidic solution. In an alternative embodiment, spent leach liquor or a combination of fresh acidic solution and spent leach liquor, can be used.
In another embodiment, step (a) of the subject method can be performed in the presence of ferrous and/or feme ions to promote dissolution of the iron mineralization. Ferrous ions will generally be present in recirculated process plant solutions
If desired, iron may be removed from the leachant solution using standard techniques, such as precipitation. The purpose is to remove soluble iron from any process solutions. Solvent extraction, ion exchange, reverse osmosis or other techniques can also be used to remove soluble iron.
The leach time for this embodiment is generally relatively long, and typically is in the range of from about 50 to about 120 hours. Preferably, leach time is from about 60 to about 100 hours. However, the operating conditions are much milder than conventional autoclave techniques, leading to large capital and operating cost advantages. Sulfuπc acid and alkali metal halides are easier to handle than the hydrochloric acid used m the Zoumei Jm et al. process refeπed to above.
The present invention provides m another separate embodiment, a method for recovenng titanium from a titanium and iron-containing mineral feed, the method including the steps of (a) solubihzmg titanium and iron by leachmg the feed with an acidic solution in the presence of an alkali metal halide and a source of activated carbon;
(b) selectively precipitating titanium oxιde(s), and
(c) recoveπng titanium oxιde(s) from the leach residue. The present invention provides in another separate embodiment a method for recoveπng titanium from a mineral feed compπsmg llmemte (FeTιO3), the method including the steps of
: (a) solubihzmg titanium and iron in the llmemte by leachmg the llmemte with an acidic solution m the presence of an alkali metal halide and a source of activated carbon; (b) selectively precipitating titanium oxιde(s), and
(c) recovering titanium oxide from the leach residue as TιO2. The present invention provides m another separate embodiment a method for recovenng titanium oxιde(s) from a mineral feed comprising llmemte (FeTιO3), the method including the steps of : (a) leachmg the llmemte with an acidic solution at a temperature m the range of from about 80 to 120°C m the presence of an alkali metal halide for a predetermined time, the leach solution containing up to about 60% by weight solids to produce a leachant solution containing iron and titanium ions; (b) separating the iron from the titanium in the leachant solution; and (c) recovering the separated titanium as Tι02.
As mentioned above, maintaining the titanium in solution rather than allowing it to report to the residue as a precipitate has been observed to further enhance the likelihood of the titanium being recovered as a pure product. Where most of the titanium reports to the residue, other materials that may be found in proximity with the llmemte mineral including chromite, lime, magnesia, silica or silicates, manganese, alumina, vanadium, phosphate and zirconium will also tend to remain m the residue along with undissolved iron. The presence of such materials is likely to dilute the puπty of titanium recoverable from the residue.
Depending on the metals content of the leach solution, a typical reaction time for step
(a) of this embodiment is up to about an hour. Preferably, the reaction time of step (a) is up to about half an hour. More preferably, the reaction time is m the range of from about 5 to about
15 minutes. It has been observed that titanium solubility reaches a peak during reaction times of approximately that length.
Optionally, step (a) above may be repeated to solubilize unleached titanium in the residue obtained following step (a) in order to obtain cumulative maximum solubility of titanium. Fresh acidic solution and alkali metal halide can be used when step (a) is repeated. Step (a) may m one embodiment comprise a type of countercurrent leach circuit.
The acidic solution in this embodiment can be supplemented with hydrochlonc acid in one or more steps of a repeated leach sequence to assist m enhancing the titanium solubility profile.
In another separate embodiment the present invention provides a method for recoveπng titanium from a titanium and iron-containmg mineral feed, the method including the steps of:
(a) contacting the feed mateπal with an halide acid solution or an acid — alkali halide solution for a period of time sufficient to solubilize the titanium but insufficient to allow the titanium to appreciably precipitate;
(b) selectively precipitating titanium oxιde(s); and
(c) recovering titanium oxιde(s).
The halide acid used in step (a) can be, for example, hydrochloric acid or hydrobromic acid. The concentration of halide acid used can be in the range of from about 150 to about 350 grams per liter acid.
Any precipitated titanium reporting to the leach residue of this embodiment may be recovered m subsequent leachmg operations.
The present invention provides m another separate embodiment a method for recoveπng titanium from a feed compπsmg finely ground llmemte (FeTι03), the method including the steps of
(a) leachmg the llmemte with an acidic solution containing sulfunc acid at a temperature of about 100°C in the presence of an alkali metal halide selected from the group consisting of NaCI, KCl and KBr and in the presence of a source of activated carbon for up to about half an hour to produce a leachant solution containing iron and titanium ions, the ratio of alkali metal halide to llmemte in the feed being about 1.2.1 ; and the ratio of activated carbon to llmemte in the feed being about 0.01 :1, the solids content of the leach solution being up to about 60%) by weight;
(b) repeating step (a); (c) separating at least some of the pregnant solution from the leach residue;
(d) selectively precipitating the titanium oxιde(s) from the pregnant solution; and
(e) recoveπng the titanium oxide as TιO2.
In a particularly preferred embodiment, the present invention provides multistage leaching of iron and titanium from an iron and titamum-bearmg mineral feed, the method compπsing the following steps. (a) contacting the feed mateπal with an acid — alkali halide solution for a penod of time sufficient to solubilize the titanium but not so long as to allow the titanium to appreciably precipitate;
(b) separating the pulp from the leach liquor; (c) contacting the pulp with fresh leach liquor and repeating steps (a) and (b) until all economically feasible titanium is leached; and (d) selectively recoveπng the titanium and iron m separate stages from the leach solutions by precipitation, solvent extraction or other means. The conditions of step (a) can involve percent solids on a weight/weight basis of between about 1 percent and about 60 percent. The typical percent solids are in the range of from about 10% to 40% The solids may be ground to fine size to facilitate leachmg, typically so that the feed passes a 73 micron sieve The acid used is most typically sulfunc acid. The acid concentration can range from about 20 to about 300 grams per liter acid. Most typically the acid concentration ranges from about 150 to 230 gram per liter. The alkali halide can be any alkali halide. Preferably, the alkali halide is NaCI, KCl,
NaBr, or KBr. The concentration of alkali halide can range from about 50 grams per liter to about 400 grams per liter. Preferably, the alkali halide concentration is about 100 to about 200 grams per liter.
The leaching is most typically carried out at about room pressure The temperature can be between about 40°C and about 110°C at room pressure Preferably, leachmg temperature is between about 90 °C and about 105°C. Leaching at room pressure will typically be performed m a leach vessel with a condenser to limit the loss of halide acid generated m the leach solution The titanium is allowed to reach a concentration as high as possible before it begins to re- precipitate onto the leach feed matenal. This is typically slightly over four (4) grams of titanium per liter of solution. The leach time to accomplish this solubihzation will depend on the vaπous aforementioned parameters but will usually range from about 10 minutes to 1 hour
The method of solid — liquid separation in step (b) can be any method that makes a good separation of the solids from the leach liquor in a relatively short time. These include methods such as cyclones, filters, centnfuges, magnetic separators, and settlers. The list is not meant to exclude any unnamed method.
The fresh leach liquor in step (c) can be leach liquor from which the titanium content has been reduced or eliminated. The iron content of liquor should be controlled so that no precipitation of an iron compound occurs dunng the leachmg
The titanium can be totally or partially removed from the leach liquor in step (d) by the method that makes the most economic sense for any given plant. The methodology available includes, but is not limited to, precipitation by seeding or pH adjustment, crystallization, solvent extraction, and ion exchange.
The iron can be removed in a similar fashion in a step before or after the titanium recovery. Titanium and iron are recovered as separate products, in separate stages. The titanium would be recovered as a titanium salt, most typically TιO2. The iron would most typically be recovered as an iron salt such as ferrous chloπde or sulphate.
In addition to titanium and iron leachmg, the present invention also concerns methods for the recovery of other minerals, such as nickel, cobalt, copper, molybdenum, lead, zmc, gold or silver from ore, soil, concentrate, slag or residue. In one embodiment, a method is provided for the dissolution of nickel and cobalt from a nickel, cobalt and iron-containing mineral feed, the method compnsmg solubihzmg the nickel, cobalt and iron m the feed by leachmg the feed with an acidic solution. In a further embodiment, an alkali metal salt can be added to the leach solution to improve recovery. The alkali metal salt can be for example, an alkali metal halide, alkali metal nitπte, alkali metal nitrate, alkali metal sulfite or alkali metal thiomte. The metal halide can be, for example, NaCI, KCl, NaBr or KBr, or mixtures of one or more of these. The metal sulfites can be, for example, sodium sulfite, sodium metabisulfite, sodium bisulfite, sodium dithionite, or other alkali metal or ammonium sulfite, metabisulfite, bisulfite or dithionite. The ordinarily skilled artisan, having the benefit of the teachings disclosed herein, can readily determine those alkali metal salts that can be used in conjunction with the particular acid solution used m the solubihzation step of the process. In another embodiment, the method of the invention can be conducted at above ambient temperatures and at or above atmospheric pressures prior to metal extraction by precipitation, solvent extraction or other means
Where the metals of interest are nickel and cobalt, the nickel and cobalt-contammg mineral feed is typically post beneficiation by comminution and thickening products. A typical example of nickel and cobalt mineralization is a latente ore. Alternatively, the feed may compnse a bulk latente concentrate.
One embodiment of the present method provides for recovering nickel and cobalt from a mineral feed comprising lateπte, the method including the step of solubihzmg nickel and cobalt and iron in the lateπte by leachmg the latente with an acidic solution in the presence of an alkali metal halide at a temperature not exceeding about 150°C at normal pressures prior to nickel and cobalt extraction by established precipitation, solvent extraction or other means.
Preferably, the acidic solution contains sulfuric acid. The sulfuric acid used in the leachmg step is typically at a concentration in the range of from about 20 grams per liter to about 500 grams per liter. In a preferred embodiment, the concentration of sulfuπc acid is in the range of from about 150 grams per liter to about 250 grams per liter. Preferably, the concentration of sulfunc acid is about 200 grams per liter. Other acids contemplated for use in the present invention include halide acids, for example, hydrochlonc acid or hydrobromic acid. The typical concentration of halide acid used is in the range from about 50 to about 350 grams per liter acid. The process is typically earned out m the presence of an alkali metal halide at a ratio of alkali metal halide to lateπte in the feed m the range of from about 0.05:1 to about 4:1.
Preferably, the ratio is about 0J : 1 , and most preferably about 0.2: 1.
In any of the descπbed embodiments of the invention, including those methods directed towards leaching of titanium and non-titanium transition elements from a mineral feed, the alkali metal salt may be added directly to the leach solution. Alternatively, the alkali metal salt is combined with the feed pπor to introduction of the leachmg solution. In this case, the feed may be subjected to a boildown pre-treatment (i.e., by boiling down to approximate dryness) in the presence of the alkali metal salt whereby the feed (e g , latente) surfaces are coated with alkali metal salt pπor to leachmg. In another alternative embodiment the solution of alkali salt may be sprayed on a heap of latentic ore and allowed to evaporate. Further a combination of the foregoing may be adopted. Namely, a proportion of the alkali metal salt is combined with the feed pπor to solubihzation and a proportion of the alkali metal salt is added directly to the leach solution. It is particularly preferred to concurrently remove some of the pregnant solution from the leach residue to permit separation of the nickel and cobalt to take place away from the leach residue. Typically, the process is carried out at a temperature in the range of from about 80 °C to about 120°C. Preferably, the temperature is m the range of from about 90 °C to about 110°C A typical operating temperature for the process is about 100°C.
The leach solution in the subject process preferably has a solids content of up to about 60% by weight. Preferably, the leach solution has a solids content of from about 10 to 40%. To facilitate rapid leach g, the feed can be ground into smaller particles. It is preferred that the feed be subjected to fine gnndmg. Preferably, the majority of particles in the feed are capable of passing through 75 micron sieve. Typically, at least 75% of the particles m the feed are of a size that can pass through 75 micron sieve openings.
Methods according to the present invention may be earned out at or above atmosphenc pressure. When elevated pressures are used, the typical elevated pressures and temperatures at which methods according to the invention may be performed are in the range of from about 1 bar to about 30 bar. Preferably, pressures are m the range of from about 1 bar to about 5 bar and temperatures range from about 100°C to about 235 °C. Preferably m the range of from about 100°C to about 150°C. The methods descπbed in the embodiments of the present invention do not conflict with known autoclave technology as the present invention involves the use of alkali metal hahdes in combination with sulfunc acid whereas known autoclave technology utilizes pure acid or ammomacal solutions to leach the nickel and cobalt from latentic feed ores.
The leach residue produced by the present process may be subjected to further leachmg to solubilize undissolved iron and/or nickel and cobalt m the residue. The further leachmg can be performed using fresh acidic solution. In an alternative embodiment, spent leach liquor, or a combination of fresh acidic solution and spent leach liquor, may be used in the process.
Additionally, the process may be performed m the presence of ferrous and/or feme ions to promote dissolution of the iron mineralization. Ferrous ions will generally be present in recirculated process plant solutions. Depending on the metals content of the leach solution, a typical reaction time for the process of this embodiment is up to about six hours. Preferably, the reaction time is up to about two hours. More preferably, the reaction time is in the range of from about 15 minutes to about 3 hours. It has been observed that nickel and cobalt solubility reaches a peak dunng reaction times of approximately that length. A person of ordinary skill in the art can vary leach time so as to leach less of an unwanted species such as manganese or iron at the expense of some cobalt and nickel recovery.
The process above may be repeated to solubilize unleached nickel or cobalt m the residue m order to obtain cumulative maximum solubility of nickel and cobalt. Fresh acidic solution and alkali metal halide may be used when the process is repeated. The process may in one embodiment comprise a type of countercurrent leach circuit.
The acidic solution may in this embodiment be supplemented with hydrochloric acid in one or more steps of a repeated leach sequence to assist in enhancing the nickel or cobalt solubility profile.
In another embodiment of the present invention, a metal halide salt may be used either to precondition an aqueous slurry or it may be sprayed onto the feed material and allowed to evaporate prior to contacting with sulfunc acid.
Upon contact with sulfuπc acid the resultant slurry is permitted to leach for a short time
(typically less than about fifteen minutes) but most preferably about five minutes or less. The liquid is then separated and sent for cobalt recovery. This flash leachmg process utilizes the selective nature of the leach to achieve a cobalt nch solution containing only minor amounts of nickel, manganese, iron, etc.
The residue from the flash leach is subsequently leached with the metal halide sulfunc acid mixture for longer periods of time to solubilize the nickel and any remaining cobalt
In another embodiment super alloy scrap and other recycled metal alloys may be leached by treating with a halide salt of the metal and sulfunc acid. The concentrations of the metal halide salt and the sulfunc acid will be dependent upon the specific scrap mixture. This embodiment can be utilized to selectively leach specific metals or to place all the metals into solution. This embodiment may also be used to solubilize radio-nucleosides of such metal as nickel from a radiated scrap. Oxygen or other oxidizing gasses such as chloπne can be added to the system to oxidize the metal.
For some oxide ores containing minerals that contain multivalent transition metals such as Co and Mn m an high oxidation state species, the alkali metal halide may be substituted with a sulfur-based reducing chemical. For example, sodium sulfite, sodium metabisulfite, sodium bisulfite, sodium dithionite, or other alkali metal or ammonium sulfite, metabisulfite. bisulfite or dithionite can be used in place of the alkali metal halide. These sulfur based reducing chemicals will facilitate the reduction of the transition metal, opening the ore up to attack by the sulfuπc acid. The metal of economic interest need not always be the one reduced. Alkali metal nitrates or nitntes may be used with sulfunc acid to leach most metals. These techniques may be used to leach metals from sulfide minerals or from scrap, residue, slags, concentrates, or soils. In another embodiment the process utilizing a metal halide salt and sulfunc acid may be used, with minor modifications, in currently existing counter current decantation (CCD) circuits. Such an embodiment would utilize fresh feed matenal to achieve neutralization to a pH adequate to retain iron in solution. After a liquid-solid separation has been effected, the resultant leach liquor may be further neutralized to precipitate iron as a hydroxide m the presence of a binding material. The iron precipitate may then be partially dned and pelletised to produce pig iron feed stocks.
The method of solid-liquid separation can be any method that produces a good separation of the solids from the leach liquor in a relatively short time. These include, but are not limited to, methods such as cyclones, filters, centrifuges, magnetic separators, and settlers. The nickel or cobalt can be totally or partially removed from the leach liquor by the method that makes the most economic sense for any given plant. The methodology available includes, but is not limited to, precipitation by seeding or pH adjustment, crystallization, solvent extraction, and ion exchange.
The subject invention also concerns methods for recovering multiple metal or metal oxides in separate solutions. Mineral species of economic value are often associated with species that consume the chemical reagents that are used to leach them. Sometimes even though the consuming species is of economic value, the overall leach becomes uneconomic. The most common example of this is copper-gold ores. Following are examples which illustrate procedures for practicing the invention. These examples should not be construed as limiting. All percentages are by weight and all solvent mixture proportions are by volume unless otherwise noted.
Example 1 — Leachmg of Titanium and Iron from llmemte
Kinetics expenments on the leaching of titanium and iron from llmemte shows that both are leached early on and that the titanium then precipitates and slows the iron leachmg. An experiment with 100 grams of llmemte, ground to -200 mesh was conducted. The tests were conducted with 1000 grams 200 gram per liter (g/1) sulfuπc acid with 120 g/1 NaCI solution. 100 grams activated carbon were added and the solution heated to 100°C. The Fe and Ti concentrations were monitored duπng the course of the 96 hour leach. The results are presented in Figure 1. The results present a mechanism of initial Ti leachmg into the liquor with subsequent hydroxylation and subsequent precipitation. While this occurs it slows leachmg of the iron. The Ti appears to be leached withm one hour. In a separate but analogous expeπment a 100 gram quantity of llmemte with a head assay of 34.0%> Fe and 27.0% Ti, and particle size such that 100% of the particles pass through a 75 micron screen, was leached for 72 hours at 100° C m 1 liter of 200 gram per liter H2SO4 — 120 gram per liter alkali metal halide solution. A 100 gram quantity of activated carbon was also present in the leach solution. The leach liquor was monitored periodically for Ti and Fe content. The results of the experiment are shown in Table 1. Titanium is dissolved then observed to subsequently precipitate. The final assay of the 57 4 gram residue showed that it contained only 0.67% Fe and 46.6% Ti. Thus, 98.9% of the iron had been extracted into the solution while 99.7% of the titanium remained in the residue. The expeπment indicates that due to the initial solubihzation of Ti, both Ti and Fe can ideally be extracted from llmemte by repeated short duration leaches.
Figure imgf000017_0001
Example 2 — Consecutive One-Hour Leachmg of Titanium and Iron from llmemte
Using the data in Example 1 allowed the development of a new leach procedure for llmemte. The procedure compπses leachmg llmemte for one hour or less and then contacting it with fresh leach solution. In this manner both the iron and titanium is leached together. This was tested using the same conditions as m the 96-hour test. The results of four (4) consecutive one-hour leaches on the same ore sample are shown m Figure 2. As can be seen, approximately the same amount of iron and llmemte was leached in each step. The ordinarily skilled artisan, having the benefit of the teachings descπbed herein, can determine the proper reagent concentration, temperature, particle size of the ore, whether to include carbon and its form (e.g., activated carbon or graphite), or atmosphenc pressure (typically < 3 atmospheres) that is optimum for a particular ore. The technique of separating the Ti as Tι02 with short leach times followed by precipitation of TιO2 is also applicable to other leach systems such as the hydrochlonc acid leach system.
The following two expenments further demonstrate methodology for leachmg both the titanium and iron from llmemte in a multistage fashion:
Experiment A comprises a leach solution of 60 grams alkali metal halide, 100 grams H2,SO4, and 350 grams of H2O heated to 100CC m Erlenmeyer flasks on a stirπng hotplate to which is added 50 grams of minus 75 microns particle size llmemte resulting m a 9% pulp density.
Experiment B comprises a leach solution of 60 grams alkali metal halide, 100 grams H2S04, and 350 grams of H20 heated to 100°C in Erlenmeyer flasks on a stirπng hotplate to which is added 100 grams of minus 75 microns particle size ilmemte resulting m a 16% pulp density.
The ilmemte had an assay head of 30% titanium and 34% iron. The following procedure steps are applied separately to Expeπment A and Expeπment B:
Step 1. A condenser is placed on the Erlenmeyer containing the slurry compnsmg the prescπbed solution and ilmemte feed; Step 2. The slurry is stirred vigorously with a magnetic stirrer for 30 minutes with the temperature maintained at 100°C; Step 3. The Erlenmeyer and contents are cooled for a couple of minutes in a room temperature water bath; Step 4. The Erlenmeyer solution is decanted into a centπfuge rube and centrifuged at 4,000 rpm for 5 minutes; Step 5. The liquor in the ccntnfugc tube is decanted and separated from the solids into a sample bottle, volume and weight determined and retained for further test work including analysis; Step 6. The remaining solids in the centrifuge tube are weighed and then washed, with 510 grams of fresh leach solution, back into the residue remaining in the Erlenmeyer after Step 4, Step 7. The reconstituted slurry is stirred and the slurry temperature increased to
100°C; Step 8. The procedure is continued by repeating Steps 1 through 7 inclusive, a total of seven times and thus equating to a total leach duration of 4 hours; Step 9. The post centπfugmg liquors collected at each repetition of Step 5 are individually subsampled and analysed for titanium and iron;
Step 10. Calculations are conducted to determine titanium and iron contents of both solids and liquors and comparisons made with the respective elemental assay values of the ilmemte ore feed;
Step 11. The individual liquors remaining after the subsamplmg conducted in Step 9 are combined in a flask and subsampled and analysed for titanium and iron; Step 12. The titanium can be totally or partially removed from the leach liquor by the method that makes the most economic sense for any given plant. The methodologies available include, but are not limited to, precipitation by seeding or pH adjustment, crystallization, solvent extraction, and ion exchange.
The results of the experiments are shown in Table 2 and Figure 3. For both levels of percent solids the trend is for roughly a constant amount of titanium to be extracted at each step.
Figure imgf000020_0001
Example 3 — Effect of Alkali Metal Halide Experiments were conducted to evaluate the effect of an alkali metal halide on the recovery of iron from ore using the sulfunc acid process of the subject invention The results are shown in Figure 4. In this case, the salt which was used was NaCI at 0%, 5%, 15% and 25%
(w/w). These tests were performed using 200 gram per liter sulfuric acid solution and no activated carbon at 100°C on unground ore. The addition of salt speeds the reaction rate.
However, at around 15 to 20% salt (150 to 200 grams per liter), NaCI appears to become counterproductive. The total percentage of iron leached actually falls after 15% NaCI is reached.
In a separate but analogous experiment to further demonstrate the effect of an alkali metal halide on the leaching of iron out of ilmenite, the samples of 100 grams of ilmenite feed were leached with 200 grams of sulfuric acid, 700 grams of water at 100° C and varying amounts of alkali metal halide for 72 hours. The amounts of alkali metal halide were 0, 50, 150 and 250 grams representing 0, 5, 15, and 25%> (w/w) alkali metal halide solutions. The leach rate of the iron was observed. The results are shown in Table 3.
Figure imgf000022_0001
Example 4 — Effect of Gnndmg of Ore In accordance with the subject invention, grinding of the ore can be used to increase the reaction rate of leaching iron from ilmenite. This is shown in Figure 5 and Table 4. Both tests were performed using a 100 gram quantity of ilmenite placed in one liter of 200 grams per liter sulfuric acid and 150 grams per liter alkali metal halide solution heated to 100°C. The experiments were conducted on two samples of the same ilmenite feed. One experiment used course ilmenite (100% retained on a 75 micron screen) and the other experiment used fine ilmenite (100% passing through a 75 micron screen). The slurry was vigorously stirred for 72 hours and the iron concentration periodically monitored. The ground ore (finer particle sized samples) had faster early and late leach kinetics than the unground ore (coarser particle sized sample). The kinetics of the ore during the 5 to 25 hour time period was similar in both cases.
Figure imgf000023_0001
Example 5 — Addition of Carbon Source
The addition of a carbon source in the form of activated carbon or graphite speeds the kinetics of the leaching reaction. The ratios of carbon to ore tried were 1 :2 and 1: 1. The results of tests carried out at 100° C with 120 grams per liter salt and 200 grams per liter sulfuric acid are shown in Figure 6. The most cost effective carbon to ore ratio will depend on final leach conditions. A person skilled in the art, having the benefit of the current disclosure, can identify the optimal carbon to ore ratio for a particular process
In a separate but analogous expeπment to demonstrate the effect of the addition of a carbon source to the leachmg of iron from ilmenite to leave a TιO2 concentrate residue, a 150 gram per liter alkali metal halide solution was used, as opposed to the 120 grams per liter salt used in the aforementioned expenment. The experiment used ilmenite having a particle size wherein 100% of particles passed through a 75 micron screen. A varying amount of coconut shell activated carbon was placed m each container. The same carbon to sample ratios were evaluated as in the aforementioned expeπment. The amounts were 0, 50 and 100 grams of carbon for carbon to sample ratios of 0, 1:2 and 1:1, respectively. The slurry was vigorously stirred for 72 hours and the iron concentration penodically monitored The results are shown in Table 5 The leach with the 1 1 ratio of carbon to feed material had slightly better kinetics than the other two conditions.
Figure imgf000025_0001
Example 6 — Leachmg of Copper and Nickel from Latente Ore with a Sulfuric Acid-Hahde- Carbon System
This ore has an assay head of 2.36% Co, 1.26% Ni, 11.00% Fe, 10.80% Mn. A sample of 100 grams of ground, -200 mesh ore was first treated with 200 grams of NaCI dissolved in 650 grams of water The water was evaporated on a hot plate. This procedure is a speeded up simulation of spaymg a heap of ore with a salt solution and letting it evaporate naturally. The ore-salt solids were then slurπed in 200 grams of sulfuπc acid in 700 grams of water solution. The stirred slurry was brought to 100° C on a stirnng hot plate, and then 100 grams of +65 mesh, coconut shell, activated carbon was added. The test was run for 48 hours with aliquots of solution taken at 1, 4, 6, and 24 hours. The results are shown in Table 6. The extraction of cobalt was probably complete withm the first hour. The cobalt was probably precipitated by the ionic strength of the solution and not recovered until the wash solution dissolved it. After 120 hours of leachmg the ore under the same conditions except for omitting the NaCI the Co recovery was 63.9% and the Ni recovery was 58.2%.
Figure imgf000026_0001
Example 7 — Effect of Alkali Metal Halide on the Leachmg of Nickel and Cobalt from Latente 1 Expenments were conducted on two samples of 100 grams of lateπte- 1 feed, compnsing
1.0 percent nickel and 0J percent cobalt of a particle size of approximately 80%> passing 75 microns.
In the first expenment the 100 g sample was leached with 200 grams of sulfunc acid,
800 grams of water and no alkali halide at 100°C. In the second expeπment the 100 g sample was leached with 200 grams of sulfuπc acid,
800 grams of water at 100°C, and 200 grams of alkali metal halide (sodium chloride).
Each expeπment was run for a total of 6 hours with solution sampling being carried out at 0.25, 0.5, 1.0, 2.0, 4.0, and finally 6.0 hours. The results are shown m Table 7 and Figures 7 and 8. The second expeπment that utilized the halide showed significantly better results for both nickel and cobalt, and particularly cobalt.
Figure imgf000027_0001
Example 8 — Effect of Alkali Metal Halide on the Leaching of nickel and cobalt from Laterite-2
Experiments were conducted on two samples of 100 grams of laterite-2 comprising 1J percent nickel and OJ percent cobalt feed of a particle size of approximately 80% passing 75 microns. In the first experiment the 100 g sample was leached with 200 grams of sulfuric acid,
800 grams of water and no alkali halide at 100°C.
In the second experiment the 100 g sample was leached with 200 grams of sulfuric acid, 800 grams of water at 100°C, and 200 grams of alkali metal halide (sodium chloride).
Each experiment was run for a total of 6 hours with solution sampling being carried out at 0.25, 0.5, 1.0, 2.0, 4.0, and finally 6.0 hours. The results are shown in Table 8 and Figures 9 and 10. The alkali metal halide (sodium chloride) test showed significantly better results for both nickel and cobalt, particularly with regard to the speed with which full (100%) dissolution is achieved.
Figure imgf000029_0001
The above two expenments demonstrate the results for leachmg both the nickel and cobalt from two different nickel-cobalt lateπte samples The following procedure steps have been applied separately to each of the latente samples:
Step 1. A condenser is placed on the Erlenmeyer containing the slurry compπsing the prescnbed solution and lateπte feed;
Step 2. The slurry is stirred vigorously with a magnetic stirrer for the duration of the test with the temperature maintained at 100°C; Step 3. The test is sampled at predetermined times, eg., 15 minutes, 30 minutes, etc., by pipetting 10 ml of the hot slurry from the Erlenmeyer into a centnfuge tube and centrifuge at 4,000 φm for 5 minutes;
Step 4. The centnfuged timed leach solution is transferred mto a sample tube for later analysis; Step 5. 10 ml of make-up leach solution is used to wash the centnfuged residue back into the Erlenmeyer, while the Erlenmeyer continues to be agitated at 100°C on the hot plate; Step 6. At the end of the test (e.g. , 6 hours) the contents of the Erlenmeyer is poured into two centnfuge tubes, using an additional very small amount of distilled water to wash out any residue remaining on the mside lip of the Erlenmeyer, and then centrifuged; Step 7. The centnfuged liquid contents (pregnant leach solution - PLS) from both centnfuge tubes is decanted into a graduated cylinder and allow to cool; Step 8. Then having read the volume of PLS solution, approximately 20 ml is transferred into a sample tube and analysed for nickel and cobalt; Step 9. Calculations are conducted to determine nickel and cobalt contents of the liquors and compaπsons made with the respective elemental assay values of the lateπte ore feed; Step 10. Nickel and cobalt can be totally or partially removed from the leach liquor by the method that makes the most economic sense for any given plant. The methodologies available vaπously include, but are not limited to, precipitation of metallic salts by seeding, pH adjustment, or crystallisation; solvent extraction and electrowmnmg of elemental metal; and ion exchange.
Example 9 — Leachmg of Silver
This example shows the leachmg of silver from a copper refinery pilot plant's slimes The test was conducted at 100°C with 200 gram per liter sulfuπc acid and 200 gram per liter NaCI. Samples of 50 grams of slimes were leached in 500 miUiliters of solution. The leachmg was conducted for 48 hours. The results are shown m the Table 9.
Figure imgf000031_0001
ppm = parts per million
Example 10 — Leachmg of Molybdenum
A sample of molybdenum oxide ore with a head grade of 0.070% Mo was ground to minus 200 mesh and leached with agitation for 48 hours at room temperature with a solution of
100 grams per liter sulfuric acid and 100 grams per liter sodium chloride. This leach recovered
89% of the molybdenum in the sample.
Another sample of unground ore from the same mine ore was screened to select the minus 18,850 plus 833 micron (minus 3 inch plus 20 mesh) fraction. This fraction was placed in a column and the same 100 g/1 sulfuric acid, 100 g/1 sodium chloride was applied to the ore for 56 days at 0.05 gallons per minute per square foot The leach solution was recirculated continuously. This leach scheme obtained 82% recovery of the molybdenum.
Example 11 — Two Stage Leaching of Different Metals into Two Separate Leach Liquors An oxide copper ore sample, ground to minus 200 mesh, with a head grade of 0.91%>
Cu, 2.0 grams/ton Au, and 2.4 grams/ton Ag was leached with 100 gram per liter sulfuric acid for 72 hours in a stirred vessel at room temperature. The solid residue was then filtered and put into another vessel and leached for 30 hours at room temperature with a solution of 50 gram per liter of potassium bromide and enough sulfuric acid (6 ml) to adjust the pH to 1.0 with agitation. The results are shown in Table 10.
Figure imgf000031_0002
A person skilled in the art, having the benefit of the teachings of this disclosure, can adjust the acid concentration and temperature to achieve complete recovery of the copper in the first stage while maintaining excellent recovery of the Au and Ag in the second stage. The subject method can also be readily adapted to heap leaching.
It should be understood that the examples and embodiments described herein are for illustrative purposes only and that various modifications or changes in light thereof will be suggested to persons skilled in the art and are to be included within the spirit and purview of this application and the scope of the appended claims.

Claims

ClaimsWhat is claimed is:
1. A method for recoveπng titanium oxide from a titanium and iron-containmg mineral feed, said method compnsmg: (a) solubihzmg said titanium and said iron by leachmg said mineral feed with an acidic solution; (b) selectively precipitating one or more titanium oxide; and (c) recovenng said precipitated titanium oxide.
2. The method according to claim 1, wherein said titanium oxide compπses titanium dioxide.
3. The method according to claim 1, wherein said mineral feed is selected from the group consisting of ore, soil, concentrate, slag and residue.
4. The method according to claim 1, wherein said mineral feed is selected from the group consisting of ilmenite (FeTιO3), leucoxene, perovskite (CaTιO3) and titano magnetite.
5 The method according to claim 4, wherein said mineral feed is ilmenite (FeTιO3)
6. The method according to claim 1, wherein said acid solution is a halide acid.
7. The method according to claim 6, wherein said halide acid is selected from the group consisting of hydrochloric acid and hydrobromic acid.
8. The method according to claim 1, wherein said acid solution is sulfuπc acid.
9. The method according to claim 8, wherein the concentration of said sulfunc acid is between about 20 grams/liter and about 500 grams/liter.
10. The method according to claim 8, wherein the concentration of said sulfuπc acid is between about 150 grams/liter and about 250 grams/liter.
11. The method according to claim 8, wherein the concentration of said sulfunc acid is about 200 grams/liter.
12. The method according to claim 1, wherein said solubihzation step is earned out in the presence of an alkali metal halide.
13. The method according to claim 12, wherein said alkali metal halide is selected from the group consisting of NaCI, KCl, NaBr and KBr.
14. The method according to claim 13, wherein said alkah metal halide is NaCI.
15. The method according to claim 12, wherein said solubihzation step is earned out in the presence of an alkali metal halide to mineral feed ratio of from about 1 : 1 to about 2:1.
16. The method according to claim 12, wherein said alkali metal halide to mineral feed ratio is from about 1 : 1 to about 1.5:1.
17. The method according to claim 12, wherein said alkah metal halide to mineral feed ratio is about 1.2: 1.
18. The method according to claim 1, wherein said solubihzation step is carried out in the presence of a source of carbon.
19. The method according to claim 1 , wherein said precipitation step is earned out using a technique selected from the group consisting of crystal seeding, concentration, pH adjustment and temperature control.
20. The method according to claim 1, further compπsmg recovering said iron.
21. The method according to claim 1 , wherein said method is earned out at or above atmospheric pressure.
22. The method according to claim 1, wherein said mineral feed is ground into smaller particles.
23. The method according to claim 22, wherein said said particles are capable of passing through a 75 micron sieve after gnndmg.
24. The method according to claiml, wherein said solubihzation step is earned out at a temperature of about 80 °C to about 120°C.
25. The method according to claim 1, wherein said solubihzation step is earned out at a temperature of about 90°C to about 110°C.
26. The method according to claim 1 , wherein said solubihzation step is earned out at a temperature of about 100°C.
27. The method according to claim 1, wherein said solubihzation step is carried out for a duration of from about 50 hours to about 120 hours.
28. The method according to claim 1, wherein said solubihzation step is earned out for a duration of from about 60 hours to about 100 hours.
29. The method according to claim 12, wherein said alkali metal halide is added as a solution to said mineral feed prior to the solubihzation step and said alkali metal halide solution is allowed to evaporate or is boiled down to dryness.
30. A multi-stage method for recoveπng titanium and iron from a titanium and lron- bearing mineral feed, said method comprising: (a) contacting said feed material with an acid solution for a penod of time sufficient to solubilize said titanium but insufficient to allow said titanium to appreciably precipitate, wherein a pulp and a leach liquor are formed; (b) separating said pulp from said leach liquor; (c) contacting said pulp with fresh leach liquor (d) repeating steps (a) and (b) until all economically feasible titanium is leached; and (e) recoveπng, separately, said titanium and said iron from said leach liquors.
31. The method according to claim 30, wherein step (a) involves solids of between about 1 percent and about 60 percent on a weight/weight basis.
32. The method according to claim 31 , wherein said solids are of a particulate size such that said solids are capable of passing through a 75 micron sieve.
33. The method according to claim 30, wherein said acid is a halide acid.
34. The method according to claim 33 wherein said halide acid is selected from the group consisting of hydrochloπc acid and hydrobromic acid.
35. The method according to claim 30, wherein said acid is sulfunc acid.
36. The method according to claim 30, wherein step (a) is earned out in the presence of an alkali metal halide.
37. The method according to claim 36, wherein said alkali metal halide is selected from the group consisting of NaCI, KCl, NaBr and KBr.
38. The method according to claim 37, wherein said alkali metal halide is NaCI
39. The method according to claim 30, wherein said solubihzation step is earned out at a temperature of about 40°C to about 110°C.
40. The method according to claim 30, wherein said solubihzation step is earned out at a temperature of about 90°C to about 105 °C.
41. A method for recovering a transition metal other than titanium from a mineral feed beaπng said transition metal, said method comprising. (a) solubihzmg said transition metal by leachmg said mineral feed with an acid solution; (b) selectively precipitating one or more of said transition metal; and (c) recovering one or more of said transition metal
42 The method according to claim 41 , wherein said transition metal is selected from the group consisting of cobalt, nickel, copper, molybdenum, lead, zmc, gold and silver.
43. The method according to claim 42, wherein said one or more transition metal comprises two transition metals, wherein said two transition metals are nickel and cobalt.
44. The method according to claim 41 , wherein said mineral feed comprises laterite.
45. The method according to claim 41, wherein said acid solution is a halide acid.
46. The method according to claim 45, wherein said halide acid is selected from the group consisting of hydrochloric acid and hydrobromic acid.
47. The method according to claim 41, wherein said acid solution comprises sulfuric acid.
48. The method according to claim 47, wherein the concentration of said sulfuric acid is between about 20 grams/liter and about 500 grams/liter.
49. The method according to claim 47, wherein the concentration of said sulfuric acid is between about 150 grams/liter and about 250 grams/liter.
50. The method according to claim 47, wherein the concentration of said sulfuric acid is about 200 grams/liter.
51. The method according to claim 41 , wherein said solubilization step is carried out in the presence of an alkali metal salt.
52. The method according to claim 51 , wherein said alkali metal salt is selected from the group consisting of alkali metal halide, alkali metal nitrate, alkali metal nitrite, alkali metal sulfite and alkali metal thionite.
53. The method according to claim 51, wherein said alkali metal salt is an alkali metal halide selected from the group consisting of NaCI, KCl, NaBr and KBr.
54. The method according to claim 53, wherein said alkali metal halide is NaCI.
55. The method according to claim 51 , wherein said alkali metal salt is selected from the group consisting of sodium sulfite, sodium metabisulfite, sodium bisulfite, sodium dithionite, ammonium sulfite, metabisulfite, bisulfite and dithionite.
56. The method according to claim 51 , wherein said solubihzation step is earned out in the presence of an alkali metal salt to mineral feed ratio of from about 1 : 1 to about 2:1.
57. The method according to claim 51 , wherein said alkali metal halide to mineral feed ratio is from about 1 : 1 to about 1.5: 1.
58. The method according to claim 51, wherein said alkali metal halide to mineral feed ratio is about 1.2: 1.
59. The method according to claim 41 , wherein said solubihzation step is earned out in the presence of a source of carbon.
60. The method according to claim 41 , wherein said precipitation step is earned out using a technique selected from the group consisting of crystal seeding, concentration, pH adjustment and temperature control.
61. The method according to claim 41, wherein said method is earned out at or above atmospheric pressure.
62. The method according to claim 41, wherein said mineral feed is selected from the group consisting of ore, soil, concentrate, slag and residue.
63. The method according to claim 41, wherein said mineral feed is ground into smaller particles.
64. The method according to claim 63, wherein said particles are capable of passing through a 75 micron sieve after gπndmg.
65. The method according to claim 41, wherein said solubihzation step is earned out at a temperature of about 80°C to about 120CC.
66. The method according to claim 41, wherein said solubihzation step is earned out at a temperature of about 90°C to about 110°C.
67. The method according to claim 41, wherein said solubihzation step is earned out at a temperature of about 100 ° C .
68. The method according to claim 41 , wherein said solubihzation step is earned out for a duration of from about 50 hours to about 120 hours.
69. The method according to claim 41 , wherein said solubihzation step is earned out for a duration of from about 60 hours to about 100 hours.
70. The method according to claim 51, wherein said alkali metal salt is added as a solution to said mineral feed pnor to the solubihzation step and said alkah metal salt solution is allowed to evaporate or is boiled down to dryness.
71. A method for consecutive recovery of multiple transition metals from a mineral feed beanng said transition metals, said method compnsmg: (a) solubihzmg a first transition metal by leachmg said mineral feed with an acid solution, whereby said mineral feed becomes separated into said first transition metal and a feed residue, said feed residue bearing at least a second transition metal; (b) selectively precipitating said first transition metal; (c) solubihzmg at least said second transition metal by leachmg said feed residue with a second acid solution additionally compnsmg an alkali metal salt; and (d) selectively precipitating at least said second transition metal.
72. The method according to claim 71, wherein said first acid solution is a halide acid.
73. The method according to claim 72, wherein said halide acid is selected from the group consisting of hydrochloric acid and hydrobromic acid.
74. The method according to claim 71, wherein said first acid solution comprises sulfuric acid.
75. The method according to claim 71, wherein said second acid solution comprises a halide acid.
76. The method according to claim 75, wherein said halide acid is selected from the group consisting of hydrochloric acid and hydrobromic acid.
77. The method according to claim 71, wherein said second acid solution comprises sulfuric acid.
78. The method according to claim 71, wherein said alkali metal salt of step (c) is selected from the group consisting of alkali metal halide, alkali metal nitrate, alkali metal nitrite, alkali metal sulfite and alkali metal thionite.
79. The method according to claim 71, wherein said alkali metal salt of step (c) is an alkali metal halide selected from the group consisting of NaCI, KCl, NaBr and KBr.
80. The method according to claim 79, wherein said alkali metal halide is KBr.
81. The method according to claim 71 , wherein said alkali metal salt of step (c) is selected from the group consisting of sodium sulfite, sodium metabisulfite, sodium bisulfite, sodium dithionite, ammonium sulfite, metabisulfite, bisulfite and dithionite.
82. The method according to claim 71, wherein said first transition metal is selected from the group consisting of cobalt, nickel, copper, molybdenum, lead, zinc, gold and silver.
83. The method according to claim 82, wherein said first transition metal is copper.
84. The method according to claim 71, wherein said second transition metal is selected from the group consisting of cobalt, nickel, copper, molybdenum, lead, zinc, gold and silver.
85. The method according to claim 84, wherein said second transition metal is gold or silver.
86. The method according to claim 71, wherein said feed residue bears a third transition metal, step (c) further comprises solubilizing said third transition metal, and step (d) further comprises selectively precipitating said third transition metal.
87. The method according to claim 86, wherein said third transition metal is selected from the group consisting of cobalt, nickel, copper, molybdenum, lead, zinc, gold and silver.
88. The method according to claim 87, wherein said third transition metal is gold or silver.
89. The method according to claim 71, wherein said mineral feed is selected from the group consisting of ore, soil, concentrate, slag and residue.
90. The method according to claim 71, wherein said solubihzation steps (a) and (c) are carried out at room temperature.
91. The method according to claim 71 , wherein each of said solubihzation steps (a) and (c) are carried out for a duration of from about 30 hours to about 120 hours.
92. The method according to claim 71, wherein said step (a) is carried out for a duration of about 72 hours and said step (c) is carried out for a duration of about 30 hours.
93. The method according to claim 71, wherein said mineral feed is ground into smaller particles.
94. The method according to claim 93, wherein said particles are capable of pass through a -200 mesh sieve.
95. The method according to claim 71, wherein said solubihzation steps (a) and (c) are carried out in the presence of a source of carbon.
PCT/US2000/004333 1999-02-19 2000-02-18 Improved methods for leaching of ores WO2000048944A1 (en)

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APAP/P/2001/002267A AP1870A (en) 1999-02-19 2000-02-18 Improved methods for leaching of ores
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AUPQ4144 1999-11-18
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US8262770B2 (en) 2007-09-18 2012-09-11 Barrick Gold Corporation Process for controlling acid in sulfide pressure oxidation processes
AU2008300273B2 (en) 2007-09-18 2012-03-22 Barrick Gold Corporation Process for recovering gold and silver from refractory ores
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MY121143A (en) 2005-12-30
CA2363031A1 (en) 2000-08-24

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