CN102203307B - Method for treating nickel laterite ore - Google Patents
Method for treating nickel laterite ore Download PDFInfo
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- CN102203307B CN102203307B CN200980143876.9A CN200980143876A CN102203307B CN 102203307 B CN102203307 B CN 102203307B CN 200980143876 A CN200980143876 A CN 200980143876A CN 102203307 B CN102203307 B CN 102203307B
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B23/00—Obtaining nickel or cobalt
- C22B23/005—Preliminary treatment of ores, e.g. by roasting or by the Krupp-Renn process
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B23/00—Obtaining nickel or cobalt
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B23/00—Obtaining nickel or cobalt
- C22B23/04—Obtaining nickel or cobalt by wet processes
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B23/00—Obtaining nickel or cobalt
- C22B23/04—Obtaining nickel or cobalt by wet processes
- C22B23/0453—Treatment or purification of solutions, e.g. obtained by leaching
- C22B23/0461—Treatment or purification of solutions, e.g. obtained by leaching by chemical methods
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B23/00—Obtaining nickel or cobalt
- C22B23/04—Obtaining nickel or cobalt by wet processes
- C22B23/0453—Treatment or purification of solutions, e.g. obtained by leaching
- C22B23/0461—Treatment or purification of solutions, e.g. obtained by leaching by chemical methods
- C22B23/0469—Treatment or purification of solutions, e.g. obtained by leaching by chemical methods by chemical substitution, e.g. by cementation
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B23/00—Obtaining nickel or cobalt
- C22B23/04—Obtaining nickel or cobalt by wet processes
- C22B23/0476—Separation of nickel from cobalt
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- Y—GENERAL 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
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- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
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Abstract
The invention relates to a method for processing laterite ores so that the metals contained in the laterites are turned into water-soluble form for the recovery of valuable metals, such as nickel and cobalt. Different types of nickel laterites are processed at the same time without being separated according to their iron and/or magnesium content. When laterites are pretreated with concentrated mineral acid so that the metals contained in the laterites react to form water-soluble salts, the silicates contained in the laterites are partially decomposed and postleaching liquid-solid separation becomes easier than it was previously. According to the method, the water vapour generated in the ore and acid reaction stage is utilised in ore drying and the unreacted mineral acid is recycled to the front end of the process.
Description
Invention field
The present invention relates to a kind of method processing laterite, so as by convert metals contained in laterite to be water miscible form to reclaim metal values, such as nickel and cobalt.Dissimilar nickel laterite can process simultaneously, and need not be isolated according to the content of their iron and/or magnesium.When using dense mineral acid to carry out pre-treatment to make metal reaction contained in laterite form water-soluble salt to laterite, silicate contained in laterite by decomposed, and rear leach in liquid-solid separation become and be more prone to than before this.According to the method, the water vapour produced in ore and sour conversion zone is used for the drying of ore, unreacted mineral acid is recycled to the front end of technique.
Background of invention
Disclose a kind of method in US Patent No. 4125588, wherein before nickel leaches, with concentrated acid, pre-treatment is carried out to nickel laterite ores.This method describe and drying is carried out to make the moisture content of ore be less than 1% to laterite ore.Be 65-100 order by dried ore grinding to grain graininess scope.Ore after pulverizing is mixed in concentrated acid with mass ratio about 1: 1.By adding water to start metal sulfate salinization reaction in the mixture containing laterite and acid, the amount adding water accounts for the ratio of the 3-40% of the quality of laterite.Metal after sulphating leaches in water.
U.S. Patent application US2006/0002835 describes a kind of method, and wherein laterite was leached by two stages.At first stage, laterite ore mixes with the vitriol oil.At second stage, ore/acid mixture, by water slurry, then leaches, and dissolves to make nickel and cobalt.The feature of the method is, is stoichiometric ratio for pretreated sulfuric acid dosage with regard to the non-ferrous metal in ore, but not is stoichiometric ratio with regard to iron.According to embodiment 5, acid/ore deposit ratio is 0.65.Acid excessive is on a small quantity favourable, and a small amount of iron is also dissolved, because this makes nickel and cobalt realize maximum dissolving.The leaching stage of the method can be carried out at 95-105 DEG C, or carries out in autoclave, and in this autoclave, temperature maximum is 150 DEG C, and pressure is equivalent to saturated vapor pressure.For improving the dissolving of cobalt, the reductive agent that some are suitable joins leaching stage, such as sulfurous gas.Owing to using the acid lower than stoichiometric ratio in the initial period of leaching process, only some laterite dissolves.The method for leaching-residence time of precipitate phase has to very long.The validity of the method requires that the soluble saprolite type laterite (saprolitic laterite) in laterite accounts for the overwhelming majority, thus makes it play a role in ferric iron areas of sedimentation.
The additive method reclaiming nickel and other metal values from laterite is described in prior art very all sidedly in U.S. Patent application US2006/0002835.
Because extract technology under normal pressure produces very tiny and in gelatinous leaching residue, this proposes to specially require for being separated and rinsing of solid and solution, brings very large challenge to the method.Although have much based on the process patent of normal pressure leaching, all do not give enough concerns for this problem.The difficulty of solid-liquid separation also certain degree depends on the kind of laterite.
Goal of the invention
The object of the invention is to eliminate produced problem in method comparatively early, and realize a kind of method leaching laterite, special feature is solid and solution separating not to cause problem.The present invention also aims to by utilizing the heat of generation in reaction and reclaiming the economy that the acid do not consumed in reaction improves laterite treatment process.
Summary of the invention
The present invention relates to a kind of nickel laterite ores that processes to promote to reclaim the method for nickel and cobalt and liquid-solid separation.According to present method,
A) steam from processing section is below utilized to carry out drying to the laterite ore pulverized,
B) dried ore is dry grinded,
C) fine ore is sent to the acid treatment mixing section using dense mineral acid to carry out, and is wherein mixed into equably in acid by ore, and sour consumption is at least stoichiometric ratio with regard to metal contained in ore,
D) the acid treatment conversion zone carried out under the mixture of ore and acid being dropped in environmental stress and the temperature between 150 DEG C and the boiling point of acid, thus described acid and ore are reacted each other; The water vapour that reaction generates is recovered and is cycled to used in the drying of ore,
E) sent to by the laterite ore after acid treatment and carry out acid recovery, wherein in the reaction of acid treatment mixing section, remaining unreacted acid is reclaimed by evaporation, and acid is subsequently cooled and is circulated back to described mixing section,
F) laterite ore after the acid treatment leaving acid recovery is delivered to metal extracting stage, this section utilizes water to carry out.
Feature according to method of the present invention is, the particle diameter of the laterite ore after pulverizing be less than 10mm account for 90%.In the dry grinding stage, it is favourable that particle diameter ore being milled to 90% is less than 500 μm, and preferably 90% is less than 150 μm.
According to the method, at the temperature of metal with the not yet large quantitative response of acid, acid and laterite are mixed into uniform mixture.
In the acid treatment of foundation the method, this acid treatment comprises mixing section and the conversion zone of acid and ore, and the metal in laterite ore forms the water-soluble salt of this mineral acid.Mineral acid is following at least one: sulfuric acid, nitric acid, or hydrochloric acid, or their at least both mixture.The concentration of mineral acid is preferably 70-98wt%.When mineral acid is sulfuric acid, conversion zone is sulphating process.
Feature according to method of the present invention is, the input acid of conversion zone and the mixture of ore is heated this section of front portion.In the method, the feature of acid recovery section is evaporated from the laterite after acid treatment at unreacted acid.The boiling point evaporated by such as the laterite after acid treatment being heated at ambient pressure acid carries out or uses negative pressure to carry out.
According to one embodiment of the invention, the slurry leaving water logging section is delivered to neutralizing zone, neutralizing zone by and slurry iron is precipitated.
According to another embodiment of the invention, the slurry that water logging goes out middle formation is directly sent to ion exchange treatment, and its ion exchange resin is for nickel and cobalt selectively.After ion exchange treatment, waste slurry is conveyed to neutralization, to make other precipitated metals.
According to third embodiment of the invention, the slurry that water logging goes out middle formation is sent to cementation (cementation) section, at this section of metal values nickel and cobalt by iron powder cementation from solution.After to metal values cementation and Magneto separate, waste slurry is conveyed to neutralization, to make other precipitated metals.
According to fourth embodiment of the invention, neutralizing agent is dropped in the water extracting stage of the material after acid treatment, iron is precipitated with the form of oxyhydroxide, and nickel and cobalt still exist with water-soluble salt form.Neutralizing agent is preferably lime and/or milk of lime.After leaching and iron precipitation stage, with the form of oxyhydroxide, sulfide or carbonate by nickel and cobalt precipitation.
According to another embodiment of the present invention, mineral acid used is sulfuric acid, and the material after at least some acid treatment is sent to thermal treatment, and in thermal treatment, iron sulfate is decomposed into sulfur oxide and rhombohedral iron ore.Sulfur oxide is sent to sulphuric acid plant and prepares sulfuric acid, and this sulfuric acid is used for the acid treatment of laterite ore.The heat treated material that leaves containing rhombohedral iron ore and water-soluble metal salt is conveyed to water logging.
List of drawings
Fig. 1 is the schema of one embodiment of the invention, and
Fig. 2-5 is schemas of other embodiments of the present invention.
Detailed Description Of The Invention
According in method of the present invention, all dissimilar nickel laterites, such as limonite, saprolite ore deposit and nontronite, or their mixture, together by dense mineral acid (70-98%) process, with the water-soluble salt making metal contained in laterite form mineral acid.Even if pending laterite exists with several dissimilar laterite, according to method of the present invention, still do not need them separated from one another.If master metal nickel, cobalt, manganese, magnesium, aluminium, chromium and iron contained in laterite, wherein nickel and cobalt are mainly thought metal values.Mineral acid mainly refers to sulfuric acid, hydrochloric acid or nitric acid, or their mixture.Hereafter for simplicity, we mention sulfuric acid, but the method is equally applicable to other mineral acids.
Also for ease of for the purpose of simplification, employ the laterite of the process of term sulphating and sulphating, but the use limiting other acid according to method of the present invention do not meaned in these terms.
Usually, in the normal pressure leaching of laterite, liquid-solid separation is proved to be existing problems.The object of the invention is to realize a kind of method, Gu thus make liquid/stalling characteristic good, and therefore likely make the method that the method becomes economic.
This technique particularly importantly, can carry out under optimum condition by the mixing of acid and laterite.In this case, that to be laterites be main points is dry and have enough thin particle diameter.But normal enriched material rugosity is enough to the degree of grinding meeting laterite, in other words, the method does not need actual fine grinding.If particle diameter is excessively thick, acid cannot be penetrated into the inside of laterite particle, and result generates water-soluble salt, namely when the sulphating process of sulfuric acid, and will be slack-off and often cannot complete.On the other hand, due to the solution heat of acid, use moist laterite can cause the activation of sulfation reaction, obstruction laterite and acid mix by it.
Be described by schema in fig. 1 according to method of the present invention.According to the method, the particle diameter that laterite ore is crushed to 90% is less than 10mm.The ore pulverized is dry with water vapour in dryer section 1, and this water vapour produces in the sulfation reaction in conversion zone 4 subsequently.Dried laterite ore is sent to dry grinding section 2, and the particle diameter there by using such as ball mill it to be typically milled to 90% is less than 500 μm, and preferably 90% is less than 150 μm.
After dry and grinding, laterite is sent to acid treatment mixing section 3, wherein laterite ore and vitriol oil Homogeneous phase mixing each other in suitable equipment, and described equipment is such as mixing screw, drum reactor, or the reactor of other kinds.The add-on of acid is at least stoichiometric ratio with regard to the metal in laterite.The temperature of mixing section is adjusted to preferably lower than 100 DEG C, make the mixture of generation non-sclerous, and be easy to further process.Its objective is and to operate at the temperature that sulphating not yet occurs.
After mix stages, mixture is delivered to conversion zone 4, wherein first improves technological temperature by indirect heating, the reaction after this between laterite and acid starts producing heat, and technique becomes autothermal to a great extent.Conversion zone is at normal pressure and carry out at the temperature of 150-300 DEG C greatly.Such as heat can be added, to make required chemical reaction optimization by inner or indirect heating such as burner in the first section of technique.Technological temperature is maximum can be sour boiling point, that is, when selecting sulfuric acid, working temperature is lower than 339 DEG C.Laterite is usually containing a lot of crystal water, and in addition, such as, can produce considerable water in sulphating processing reaction, these water evaporate in heat-processed.The water vapour produced is recovered and is used to dry laterite, as previously mentioned.According to the method in US Patent No. 4125588, sulphating treatment stage, add water, but according to method of the present invention, do not add water at this conversion zone.Mixing section is the same with extracting stage afterwards with conversion zone, all occurs under ambient pressure, and that is system is at ambient pressure.Leave acid-treated material be solid and be actually powder, be therefore convenient to process.
Use excessive acid with regard to metal contained in laterite can improve the transformation of metal to water-soluble salt (sulphating process), and therefore improve the recovery degree of metal.In this case shortcoming is the increase in the cost of acid and neutralization.Consumed by the acid using the circulation of unreacted residual acid to reduce in technique.Residual acid evaporates in acid recovery section 5 from the laterite after acid treatment.Such as by add hot acid treatment after laterite to acid boiling point 339 DEG C evaporate.Another kind of mode uses negative pressure, makes the boiling point lowering of acid.After evaporation, acid is cooled and is condensed into liquid, and is cycled back to mixing section 3.
When laterite is such as with sulfuric acid reaction, describe by the reaction equation simplified below:
2FeOOH+3H
2SO
4＝＞Fe
2(SO
4)
3+4H
2O (1)
NiO+H
2SO
4＝＞NiSO
4+H
2O (2)
2CoOOH+3H
2SO
4＝＞Co
2(SO
4)
3+4H
2O (3)
Al
2O
3+3H
2SO
4＝＞Al
2(SO
4)
3+3H
2O (4)
MgO+H
2SO
4＝＞MgSO
4+H
2O (5)
M
2O+H
2SO
4＝＞M
2SO
4+H
2O(M＝Na，K) (6)
According to method of the present invention, utilize following phenomenon, according to this phenomenon, silicate contained in laterite can partial dehydration in concentrated acid treating processes, and occurs in this phenomenon carry out under also may being higher temperature each section.Meanwhile, the crystalline structure of silicate is with a kind of mode generating portion change being beneficial to liquid-solid separation subsequently.
According to Fig. 1, after acid treatment and acid recovery section, the solid fines of the water-soluble salt containing this acid is sent to actual extracting stage 6, is delivered in solid by water in this section.Water extracting stage carries out in atmospheric conditions, that is, at the temperature of 80-105 DEG C and environmental stress.The time length of this extracting stage depends on the composition of particle diameter and laterite ore, usually between 1-2h.
The extracting stage 6 of the laterite after acid treatment, the dissolving metal of all sulphatings contained in laterite.According to the alternatives of Fig. 1, by neutralizing zone 7 with some suitable neutralizing agent neutralization solutions, mainly with trivalent exist iron be separated from pregnant solution, to make iron precipitated, nickel salt and cobalt salt remain water miscible.Preferred neutralizing agent is Wingdale and/or milk of lime, thus iron is precipitated with the form of oxyhydroxide.Such as, if wish that iron is with the form of such as jarosite precipitation, can also use other known precipitation agent, sodium sulfate.Aluminium and most chromium and iron precipitate simultaneously, thus make them enter leaching residue.
Water extracting stage 6 and in and after 7, carry out solid and be separated 8, use conventional separation method to make solid and liquid separation, such as concentrated and/or filter, thus make the material that stays in solids, such as iron and silicate, with solution separating.The metal of all dissolvings, particularly comprise the solution (PLS) of metal values, be sent to section subsequently to do further process to this solution, the ion dissolved there is effectively separated, and is sent to next processing section (not describing in detail in the graph).As mentioned above, the structure of silicate changes in the process of laterite according to of the present invention, be separated to make solution-solids and do not cause gelation and be therefore easy to carry out.The feature of the method is, owing to leaching the character of residue, compared with after directly carrying out acidleach, the concentrated of solution can carry out with washing in relative much smaller equipment.Meanwhile, relative to currently known methods, in final refuse, the ratio of remaining metal values is also obvious much little.
From solution, reclaim required metal values such as nickel and cobalt undertaken by currently known methods, iron precipitation after by them with oxyhydroxide, sulfide or carbonate form precipitation, this does not describe in detail in the graph yet.Common chemical is used, such as hydrogen sulfide, sodium pyrosulfate, lime, magnesium hydroxide or sodium carbonate in precipitation process.
Use lime or a kind of corresponding neutralizing agent, (usual scope is 9-11) removes magnesium and manganese from solution at a high ph, thus Mg and Mn is mainly precipitated with the form of oxyhydroxide.
Fig. 2 describes the second alternative embodiment, for the treatment of the surplus materials leaving water logging section 6.Initially each section of the method is carried out as mentioned above.The feature of this alternatives is, do not carry out liquid-solid separation after the leaching, but whole slurry is delivered to ion exchange treatment 9, metal values such as nickel and cobalt is directly reclaimed according to the concept being called as " resin is in ore pulp " by ion exchange resin from slurry there.After ion exchange resin absorption, resin is sent to wash-out section 10, and carry out wash-out with acid or equivalent eluent there, the elute soln containing metal values is conveyed to further process.Resin is circulated back to ion-exchange section 9.From elute soln, reclaim metal values can such as be undertaken by sulfide precipitation or liquid-liquid extraction, be not described in detail herein.Leave the waste residue containing iron and silicate of ion-exchange section, be sent to neutralizing zone, at neutralizing zone, iron, aluminium, manganese and magnesium etc., be precipitated out from solution.
Fig. 3 shows the third alternative embodiment of the present invention, for the treatment of leaving the surplus materials that water logging goes out 6.In cementation section 11 by using the iron powder of suitable thickness by desired metal values, as nickel and cobalt, cementation, thus them are reclaimed from the slurry leaving extracting stage 6, and metal values has been stayed in iron powder.In this case, before recovery metal values, slurry does not also carry out liquid-solid separation.Iron powder is such as separated from slurry by weak magnetic separator in Magneto separate section 12.The mixture of iron powder and metal values is sent to metal values extracting stage 13, and mixture is by acidleach there, solution be sent to subsequently processing section to reclaim nickel and cobalt.The waste residue containing iron and silicate produced when cementation metal values is sent to neutralizing zone, and at this iron leg, aluminium, manganese and magnesium etc. are precipitated out from solution.
Also can modify according to Fig. 4 according to technique of the present invention disclosed in Fig. 1, wherein show, water extracting stage 6 after acid treatment also can carry out as associating extracting stage 14, there, the laterite ore after acid treatment directly can be processed by the Wingdale neutralizer adding aqueous phase aequum.Like this, the leaching of water-soluble salt in laterite, solution neutralization and the precipitation of iron that obtains all are carried out at same section.Only a solid-liquid separation section 15 is needed in this alternatives, herein, the throw out of the leaching residue, gypsum and the iron that are formed and solution separating.PLS solution containing nickel and cobalt goes to process subsequently, and such as sulfide precipitation section 16, and final solution goes to the removal section 17 of magnesium and manganese, by neutralization, metal is precipitated from solution herein.Final solution mainly water, it can be circulated back to extracting stage (not being shown specifically in the drawings).
Fig. 5 discloses another alternative embodiment of the present invention, for the further process to laterite after acid treatment.The material leaving mixing and conversion zone is wholly or partly and drops into thermal treatment section 18 after acid recovery section 5, and this section is carried out in drum furnace or fluid bed furnace.Heat energy needed for thermal treatment section can be obtained by such as coal combustion.When acid used is sulfuric acid, the iron reaction in acid treatment section laterite generates iron sulfate, and this iron sulfate is decomposed into rhombohedral iron ore and sulfur oxide in thermal treatment section.Sulfur oxide gas (SO
2and SO
3) being sent to sulphuric acid plant 19, the acid of producing from this sulphuric acid plant is recycled to acid treatment section 3.Material containing rhombohedral iron ore is conveyed to water logging and goes out 6, and herein, nickel, cobalt and magnesium are dissolved and reclaimed by above-described method.Before thermolysis, free acid residual in the laterite after acid treatment is evaporated by the method described before.The advantage of this technique is that waste material is easy to process, and reduces the consumption of acid.
Embodiment
By following examples, working method is described.Laterite ore in the present embodiment is nontronite, and its composition row in Table 1.On pretreatment, dry ore, and be crushed to particle diameter 100% and be less than 1mm.
The composition of table 1. nickel laterite
Embodiment 1.
The particle diameter with ball mill the laterite of drying being ground to 95% is less than 105 μm.Acid used is sulfuric acid.Sulfuric acid and laterite mix with mass ratio 1: 1 at mixing section.348g mixture drops into conversion zone rotary kiln, and the central point temperature of stove is 250 DEG C.Mixture is approximately 30 minutes in the residence time of stove hot-zone.The weight of rotary kiln product is 296.3g.The composition row of rotary kiln product in table 2.
The composition of table 2. rotary kiln product
In rotary kiln after process, in extracting stage, the laterite of 294.8g sulphating is leached with 1 premium on currency in titanium reactor.Leaching experiment continues 6 hours, and slurry temperature is 80 DEG C.In leaching process, the metal content row in solid in table 3.The weight leaching residue is 140.2g.To the metal productive rate row in solution in table 4.
The metal content of solid in table 3. leaching process
Table 4. is to the metal productive rate in solution
Embodiment 2
Grinding is described by the following examples for the importance of nickel productive rate in leaching process.Use the ore do not ground to test, it has been crushed to particle diameter 100% and has been less than 1mm.Sulfuric acid and laterite mix with mass ratio 1: 1 at mixing section.Conversion zone carries out in rotary kiln, in this rotary kiln, drop into 400g mixture.Stove central point temperature is 250 DEG C.Mixture is approximately 27 minutes in the residence time of stove hot-zone.The weight of rotary kiln product is 279.1g.Rotary kiln product in table 5 in column in groups.
The composition of table 5. rotary kiln product
In rotary kiln after process, in leaching stage, the laterite of 200g sulphating is leached with 1 premium on currency in 1 liter of titanium reactor.Leaching experiment continues 6 hours, and slurry temperature is 80 DEG C.In leaching process, the metal content of solid is listed in table 6.The weight leaching residue is 66.2g.To the metal productive rate row in solution in table 7.
The metal content of solid in table 6. leaching process
Table 7. is to the metal productive rate in solution
Embodiment 3
Acid and laterite mass ratio are illustrated by the following examples for the importance of nickel productive rate in leaching process.As the same manner as in Example 1, carry out another two experiments, to be acid with the ratio of laterite be difference 0.6 and 0.8kg/kg.The acid amount consumed in the sulfation reaction of this process can after water logging goes out by volumetric soiutions in remaining acid determine.Table 8 lists acid and the impact of laterite ratio for the amount of metal productive rate and unreacted acid.Show in table, when increasing acid amount corresponding to laterite weight, metal productive rate is obviously improved.In this case, unreacted acid amount also increases.Result shows, and for maximum nickel productive rate, the amount of acid used and the mass ratio of laterite should be approximately 1: 1.When the method introduced with the present invention is by sour circulation time, the great amount of cost saving relevant to the acid consumed can be realized.
Table 8. adopts different acid-laterite ratio, to the metal productive rate in solution and remnants acid amount
Claims (23)
1. process nickel laterite ores to reclaim nickel and cobalt and to promote the method for liquid-solid separation, it is characterized in that according to present method,
A) utilize from below treatment stage (d) steam to pulverize laterite ore carry out drying (1),
B) carried out to dried ore the dry grinding stage (2),
C) fine ore is sent to the acid treatment mix stages (3) using dense mineral acid, wherein ore is mixed in acid equably, acid consumption with regard to all metals contained in ore, be at least stoichiometric ratio and wherein said dense mineral acid comprise from below treatment stage (e) recovered acid
D) mixture of ore and acid is dropped into the acid treatment step of reaction (4) of only carrying out under environmental stress and the temperature between 150 DEG C and the boiling point of described acid, react each other to make described acid and ore; The water vapour generated in reaction is recovered and is cycled to used in the drying (1) of ore,
E) sent to by the laterite ore after acid treatment and carry out the acid recovery stage (5), wherein unreacted acid is reclaimed by evaporation, and this acid is subsequently cooled and is circulated back to acid treatment mix stages (3),
F) laterite ore after the acid treatment leaving acid recovery is delivered to water leaching stage (6), this water leaching stage utilizes water to carry out.
2. according to the method for claim 1, be characterised in that the particle diameter of the laterite ore of pulverizing be less than 10mm account for 90%.
3., according to the method for claim 1, be characterised in that, in the dry grinding stage (2), the particle diameter of ore grinding to 90% be less than 500 μm.
4., according to the method for claim 3, be characterised in that, in the dry grinding stage (2), the particle diameter of ore grinding to 90% be less than 150 μm.
5., according to the method for claim 1, to be characterised in that at metal not yet with at the temperature of described acid-respons, described acid and laterite are mixed into uniform mixture.
6., according to the method for claim 1, be characterised in that the metal in laterite ore forms the water-soluble salt of described mineral acid in acid treatment, this acid treatment comprises acid treatment mix stages (3) and acid treatment step of reaction (4).
7., according to the method for claim 1, be characterised in that described mineral acid is following at least one: sulfuric acid, nitric acid or hydrochloric acid, or the mixture of at least two kinds in these.
8., according to the method for claim 1, be characterised in that the concentration of described mineral acid is 70-98%.
9., according to the method for claim 1 or 7, be characterised in that described mineral acid is sulfuric acid, and acid treatment step of reaction (4) is sulphating process.
10., according to the method for claim 1, be characterised in that and the mixture of the acid and ore that drop into acid treatment step of reaction (4) is heated at the initial portion in this stage.
11. according to the method for claims 1, be characterised in that by add at ambient pressure hot acid treatment after laterite to the boiling point of described acid make unreacted acid evaporation in the acid recovery stage (5).
The method of 12. foundation claims 1, is characterised in that and utilizes negative pressure that the unreacted acid in the acid recovery stage (5) is evaporated.
The method of 13. foundation claims 1, is characterised in that and the slurry leaving water leaching stage (6) is delivered to neutralization stage (7), by neutralization described slurry, iron is precipitated here.
14. according to the method for claims 1, is characterised in that the slurry formed in water leaching stage (6) directly to send to carry out ion exchange treatment (9), and its ion exchange resin is for nickel and cobalt selectively.
The method of 15. foundation claims 14, is characterised in that and the waste residue from ion exchange treatment (9) is sent to neutralization to make other precipitated metals.
The method of 16. foundation claims 15, be characterised in that and the slurry formed in water leaching stage (6) is delivered to the cementation stage (11), in this cementation stage, metal values nickel and cobalt by iron powder cementation from solution, and are separated from slurry by Magneto separate (12).
The method of 17. foundation claims 16, is characterised in that after metal values being carried out to cementation stage (11) and Magneto separate (12), and described waste residue is conveyed to neutralization to make other precipitated metals.
The method of 18. foundation claims 1, be characterised in that in the leaching and iron precipitate phase (14) of the material after acid treatment, drop into neutralizing agent precipitates with the form of oxyhydroxide to make iron, and nickel and cobalt is still water-soluble salt.
The method of 19. foundation claims 18, is characterised in that neutralizing agent is Wingdale and/or milk of lime.
The method of 20. foundation claims 1 or 18, is characterised in that with oxyhydroxide from solution, the form of sulfide or carbonate carries out the precipitation (16) of nickel and cobalt after leaching and iron precipitate phase (14).
21., according to the method for claims 1, are characterised in that the mineral acid of use is sulfuric acid, and are sent to by the material after acid treatment at least partially and heat-treat (18), and in this thermal treatment, iron sulfate is decomposed into sulfur oxide and rhombohedral iron ore.
The method of 22. foundation claims 21, is characterised in that and sulfur oxide is delivered to sulfuric acid apparatus (19) to manufacture sulfuric acid, this sulfuric acid is used for carrying out acid treatment to laterite ore.
The method of 23. foundation claims 21, is characterised in that and the material be made up of rhombohedral iron ore and water-soluble metal salt leaving thermal treatment (18) is delivered to water leaching stage (6).
Applications Claiming Priority (3)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| FI20080603A FI121180B (en) | 2008-11-03 | 2008-11-03 | A method for treating nickel plater ore |
| FI20080603 | 2008-11-03 | ||
| PCT/FI2009/050870 WO2010061045A1 (en) | 2008-11-03 | 2009-10-29 | Method for treating nickel laterite ore |
Publications (2)
| Publication Number | Publication Date |
|---|---|
| CN102203307A CN102203307A (en) | 2011-09-28 |
| CN102203307B true CN102203307B (en) | 2015-03-04 |
Family
ID=40097248
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CN200980143876.9A Expired - Fee Related CN102203307B (en) | 2008-11-03 | 2009-10-29 | Method for treating nickel laterite ore |
Country Status (8)
| Country | Link |
|---|---|
| EP (1) | EP2350327A4 (en) |
| CN (1) | CN102203307B (en) |
| AU (1) | AU2009321543B2 (en) |
| BR (1) | BRPI0921490A2 (en) |
| CU (1) | CU23937B1 (en) |
| EA (1) | EA018749B1 (en) |
| FI (1) | FI121180B (en) |
| WO (1) | WO2010061045A1 (en) |
Families Citing this family (11)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| CN102268537B (en) * | 2011-08-15 | 2013-05-29 | 广西银亿科技矿冶有限公司 | Method for extracting cobalt and nickel from laterite-nickel ore |
| CN102286661A (en) * | 2011-08-25 | 2011-12-21 | 云南锡业集团（控股）有限责任公司 | Method for direct electrolysis of laterite nickel ore by sulfuric acid leaching |
| FI123884B (en) | 2011-11-08 | 2013-11-29 | Outotec Oyj | A process for dissolving a sulphidic metal concentrate |
| CN103361490B (en) * | 2012-03-30 | 2016-02-24 | 吉坤日矿日石金属株式会社 | The manufacture method that electricity is plumbous |
| AU2013325098B2 (en) * | 2012-09-28 | 2017-12-14 | Direct Nickel Pty Ltd | Method for the recovery of metals from nickel bearing ores and concentrates |
| FI124419B (en) * | 2013-01-29 | 2014-08-29 | Global Ecoprocess Services Oy | A method for recovering metals from oxide ores |
| JP5622061B2 (en) * | 2013-03-26 | 2014-11-12 | 住友金属鉱山株式会社 | Method for producing hematite for iron making |
| WO2015132473A1 (en) | 2014-03-06 | 2015-09-11 | Outotec (Finland) Oy | Method, arrangement and use for treating nickel ore |
| CN106987724A (en) * | 2017-04-08 | 2017-07-28 | 广西凤山县五福矿业发展有限公司 | A kind of method for solidifying silicon from the low molten aluminium slag of iron aluminium concentrate |
| CN106987723A (en) * | 2017-04-08 | 2017-07-28 | 广西凤山县五福矿业发展有限公司 | A kind of method that aluminium is reclaimed from the low molten aluminium slag of iron aluminium concentrate |
| CN113025832B (en) * | 2021-03-02 | 2022-07-15 | 重庆大学 | Nickel extraction and CO mineralization from laterite-nickel ore2Method (2) |
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|---|---|---|---|---|
| US4125588A (en) * | 1977-08-01 | 1978-11-14 | The Hanna Mining Company | Nickel and magnesia recovery from laterites by low temperature self-sulfation |
| FR2459295A1 (en) * | 1979-06-18 | 1981-01-09 | Eleusis Bauxite Mines Inc | Nickel and cobalt recovery from laterite ores - by heat treating crushed ore to oxidise iron cpds. and leaching with warm sulphuric acid |
| CN1271392A (en) * | 1997-08-15 | 2000-10-25 | 康明柯工程服务有限公司 | Chloride assisted hydrometallurgical extraction of nickel and cobalt fro m sulphide or laterite ores |
| WO2003004709A1 (en) * | 2001-07-06 | 2003-01-16 | Omg Finland Oy | Method for recovering nickel and eventually cobalt by extraction from nickel-containing laterite ore |
| WO2007074360A2 (en) * | 2005-11-28 | 2007-07-05 | Anglo Operations Limited | Leaching process in the presence of hydrochloric acid for the recovery of a value metal from an ore |
| CN101006190A (en) * | 2004-06-28 | 2007-07-25 | 斯凯资源有限公司 | Method for nickel and cobalt recovery from laterite ores by reaction with concentrated acid and water leaching |
Family Cites Families (1)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| US4298379A (en) * | 1980-01-31 | 1981-11-03 | The Hanna Mining Company | Production of high purity and high surface area magnesium oxide |
-
2008
- 2008-11-03 FI FI20080603A patent/FI121180B/en not_active IP Right Cessation
-
2009
- 2009-10-29 EA EA201100547A patent/EA018749B1/en not_active IP Right Cessation
- 2009-10-29 BR BRPI0921490A patent/BRPI0921490A2/en not_active IP Right Cessation
- 2009-10-29 EP EP09828687.5A patent/EP2350327A4/en not_active Withdrawn
- 2009-10-29 AU AU2009321543A patent/AU2009321543B2/en not_active Ceased
- 2009-10-29 WO PCT/FI2009/050870 patent/WO2010061045A1/en active Application Filing
- 2009-10-29 CN CN200980143876.9A patent/CN102203307B/en not_active Expired - Fee Related
-
2011
- 2011-04-29 CU CU20110099A patent/CU23937B1/en active IP Right Grant
Patent Citations (6)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| US4125588A (en) * | 1977-08-01 | 1978-11-14 | The Hanna Mining Company | Nickel and magnesia recovery from laterites by low temperature self-sulfation |
| FR2459295A1 (en) * | 1979-06-18 | 1981-01-09 | Eleusis Bauxite Mines Inc | Nickel and cobalt recovery from laterite ores - by heat treating crushed ore to oxidise iron cpds. and leaching with warm sulphuric acid |
| CN1271392A (en) * | 1997-08-15 | 2000-10-25 | 康明柯工程服务有限公司 | Chloride assisted hydrometallurgical extraction of nickel and cobalt fro m sulphide or laterite ores |
| WO2003004709A1 (en) * | 2001-07-06 | 2003-01-16 | Omg Finland Oy | Method for recovering nickel and eventually cobalt by extraction from nickel-containing laterite ore |
| CN101006190A (en) * | 2004-06-28 | 2007-07-25 | 斯凯资源有限公司 | Method for nickel and cobalt recovery from laterite ores by reaction with concentrated acid and water leaching |
| WO2007074360A2 (en) * | 2005-11-28 | 2007-07-05 | Anglo Operations Limited | Leaching process in the presence of hydrochloric acid for the recovery of a value metal from an ore |
Also Published As
| Publication number | Publication date |
|---|---|
| FI20080603A0 (en) | 2008-11-03 |
| EP2350327A1 (en) | 2011-08-03 |
| CU20110099A7 (en) | 2012-06-21 |
| EA201100547A1 (en) | 2011-12-30 |
| EA018749B1 (en) | 2013-10-30 |
| AU2009321543A1 (en) | 2010-06-03 |
| AU2009321543B2 (en) | 2014-09-11 |
| CU23937B1 (en) | 2013-08-29 |
| EP2350327A4 (en) | 2016-11-09 |
| WO2010061045A1 (en) | 2010-06-03 |
| FI20080603A (en) | 2010-05-04 |
| BRPI0921490A2 (en) | 2016-01-19 |
| CN102203307A (en) | 2011-09-28 |
| FI121180B (en) | 2010-08-13 |
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