AU2023263558B2 - A processing method of a laterite-nickel ore acid leaching slag - Google Patents
A processing method of a laterite-nickel ore acid leaching slag Download PDFInfo
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Abstract
The present disclosure provides a processing method of a laterite-nickel ore acid leaching slag.
The processing method includes: step S1, oxidation roasting a raw material comprising a
laterite-nickel ore acid leaching slag and a flux in a roasting kiln to obtain a high-temperature calcine
and a roasting flue gas; step S2, in an oxygen-enriched air, preliminary smelting a raw material
comprising the high-temperature calcine and a first fuel in the smelting-melting zone of the side
blown furnace to obtain a melt and a smelting zone flue gas; step S3, in a reducing atmosphere,
reduction smelting a raw material comprising the melt, a reducing agent, and a second fuel in the
reducing zone of the side blown furnace to obtain a reduced melt and a reducing zone flue gas; and
step S4, subjecting the reduced melt to slag-iron separation in the electrothermal settling zone of
the side blown furnac to obtain a smelting slag, molten iron, and an electrothermal settling zone flue
gas. The above processing method is conducive to improving the recovery rate of iron and the
product quality of molten iron. The whole process has low energy consumption, high system heat
utilization rate, and the above processing method is simple, has low process costs and high
economic benefits.
Drawings
flux leaching slag dust return a
batching system
rosing kiln
high-terperature oxygen-enriched reducing
fuel calcine air agent flue gas
side blown molten smelting waste heat, electric dust
bath smelting device flue gas bie coleto
Molten ironingot slag coal gas desulfurization
casting
up-to-standard
water quenching waste heat boiler discharge
i . smoke..
glassy state by- coal gas purification dust
products system
purifying the
coal gas
coal gas boiler
steam electric
power generation
Fig. 1
Description
Drawings
flux leaching slag dust return a
batching system
rosing kiln
high-terperature oxygen-enriched reducing fuel calcine air agent flue gas
side blown molten smelting waste heat, electric dust bath smelting device flue gas bie coleto
Molten ironingot slag coal gas desulfurization casting up-to-standard water quenching waste heat boiler discharge i . smoke.. glassy state by- coal gas purification dust products system
purifying the coal gas
coal gas boiler
steam electric power generation
Fig. 1
Technical Field
The present disclosure relates to the technical field of processing of a laterite-nickel ore acid leaching slag, in particular to a processing method of a laterite-nickel ore acid leaching slag.
Background
With the increasing demand for nickel around the world and the diminishing resources of nickel sulfide ore, obtaining nickel resources from laterite-nickel ore has gradually become a research hotspot. As the production of hydrometallurgical projects such as Ramu Project and Lygend Project, etc., has reached the design capacity, a large amount of wet leaching slag stockpiling that is difficult to be processed has been formed, causing serious environmental pollution. So far, there is no effective industrial examples for the recycling and utilization of laterite-nickel ore leaching slag, which mainly rely on tailings reservoir stockpiling and deep-sea landfill. In the existing studies, articles such as "Experimental Study on Recovery of Iron Minerals from Laterite-Nickel Ore Acid Leaching Slag", "Study on Magnetization Roasting-Weak Magnetic Separation of Iron Ore Concentrate from Laterite-Nickel Ore Pressurized Leaching Slag", and "Experimental Study on Biomass Magnetization Roasting-Magnetic Separation Recovery of Iron Ore Concentrate from Laterite-Nickel Ore Acid Leaching Slag" provide the processing of the laterite-nickel ore acid leaching slag, however, due to the fine particle size of the leaching slag and low recovery rate of iron, the selected iron ore concentrate often cannot meet the standard requirements.
Summary
The main object of the present disclosure is to provide a processing method of a laterite-nickel ore acid leaching slag, in order to solve the problems of low recovery rate of iron and poor quality of molten iron in the existing technology of recovering iron ore concentrate from laterite-nickel ore acid leaching slag.
In order to achieve the above object, according to one aspect of the present disclosure, a processing method of a laterite-nickel ore acid leaching slag is provided, and the processing method includes: step S1, oxidation roasting a raw material comprising a laterite-nickel ore acid leaching slag and a flux in a roasting kiln to obtain a high-temperature calcine and a roasting flue gas; step S2, in an oxygen-enriched air, preliminary smelting a raw material comprising the high-temperature calcine and a first fuel in the smelting-melting zone of the side blown furnace to obtain a melt and a smelting zone flue gas; step S3, in a reducing atmosphere, reduction smelting a raw material comprising the melt, a reducing agent, and a second fuel in the reducing zone of the side blown furnace to obtain a reduced melt and a reducing zone flue gas; and step S4, subjecting the reduced melt to slag-iron separation in the electrothermal settling zone of the side blown furnac to obtain a smelting slag, molten iron, and an electrothermal settling zone flue gas.
Further, in the step S3, the reducing atmosphere is a strong reducing atmosphere, which comprises carbon monoxide and carbon dioxide, and the volume ratio of carbon monoxide to the strong reducing atmosphere is 60-90:100.
Further, in the step S3, the mass of the reducing agent is 15-25% by weight of the laterite-nickel ore acid leaching slag on a dry basis, the reducing agent is preferably selected from any one or more of pulverized coal, coke powder, charcoal active powder, and graphite.
Further, in the step S3, the temperature of the reduction smelting is 1450-1550°C, and the time of the reduction smelting is preferably 0.5-2.5 hours; and the second fuel is preferably selected from any one or more of natural gas, coal gas, pulverized coal, heavy oil, pyrolysis oil, and biomass oil.
Further, in the step S1, the temperature of the oxidation roasting is 800-1200 °C, and the desulfurization rate of the oxidation roasting is preferably greater than 90%, the waste heat is preferably recovered by passing the roasting flue gas through a waste heat boiler, which is then subjected to dust collection and purification before reaching the standard for discharge.
Further, the mass ratio of the flux to the laterite-nickel ore acid leaching slag is 5-20:100, the flux is preferably limestone and/or quicklime, and the mass ratio of CaO to SiO 2 in the smelting slag is preferably controlled to be 0.5-1.5:1.
Further, in the step S2, the temperature of the preliminary smelting is 1450-1550°C, the time of the preliminary smelting is preferably 0.5-2.5 hours, and the waste heat is preferably recovered by passing the smelting zone flue gas through a waste heat boiler, which is then subjected to dust collection and purification before reaching the standard for discharge; preferably, in the step S2, the oxygen content in the oxygen-enriched air is 50-100%, and the jetting pressure of the oxygen-enriched air is preferably 0.2-0.6 MPa; and the first fuel is preferably selected from any one or more of natural gas, coal gas, pulverized coal, heavy oil, pyrolysis oil, and biomass oil.
Further, the waste heat is recovered by a waste heat boiler after the reducing zone flue gas is combined with the electrothermal settling zone flue gas, which is then subjected to dust collection and purification to obtain the coal gas, wherein 50-100% of the coal gas is sent back to the roasting kiln as a heat source of the roasting kiln, and the steam is producible from the remaining coal gas via a coal gas boiler; and preferably, the smelting zone flue gas, the reducing zone flue gas, the electrothermal settling zone flue gas, and the roasting flue gas are passed through the waste heat boiler to recover the steam produced during the process, which is combined with the steam produced by the coal gas boiler for power generation.
Further, the smoke dust obtained from the electric dust collection of the smelting zone flue gas, the reducing zone flue gas, the electrothermal settling zone flue gas, and the roasting flue gas is granulated, and which is then returned to the roasting kiln.
Further, in the step S4, the temperature for the slag-iron separation in the electrothermal settling zone is 1450-1600°C.
By applying the technical solution of the application, the laterite-nickel ore acid leaching slag and a solvent are sent to the rotary kiln for oxidation roasting, so as to desulphurize and dehydrate (where the dust returned from the smelting system can be fed together into the roasting kiln after batching). The desulfurization rate of the oxidation roasting is greater than 95%, and the chemical reactions of the sulfur elements involved in the oxidation roasting process are as follows:
Fe 2(SO 4 )3 = Fe 203+3SO2+1.502
Cr2(SO4)3 = Cr203+3SO2+1.502
A12(SO 4) 3 = A1 203+3SO2+1.502
4FeS+110 2 = 2Fe 2O3+8SO2
The high-temperature calcine sent to the smelting-melting zone of the side blown furnace (side blown molten bath smelting device) quickly melts under the large amount of heat provided by the fuel and the oxygen-enriched air, forming a melt, which is transferred to the reducing zone of the side blown furnace, and in the reducing atmosphere, the metals in the melt are deeply reduced by a reducing agent to obtain a reduced melt including metal and slag. Subsequently, the reduced melt flows into the electrothermal settling zone to further settle and separate the metal iron in the slag, after which the smelting slag is discharged through the slag notch and water quenched to obtain harmless glass slag, the molten iron is discharged through the metal notch and the ingot casting is sold, the recovery rate of iron is above 95%, and the sulfur content of the cast iron is less than 0.03%. The above processing method is conducive to improving the recovery rate of iron and the product quality of molten iron. The waste heat of flue gas generated from the whole process can be fully recovered, thereby reducing the energy consumption and improving the system heat utilization rate, and the harmless glass slag can be used as a building material, and the above processing method is simple, has low process costs and high economic benefits.
Brief Description of the Drawings
The accompanying drawings of the description, which form a part of the application, are used to provide a further understanding of the disclosure. The illustrative examples and their descriptions of the disclosure are used to explain the disclosure, and do not constitute an improper limitation thereto. In the accompanying drawings:
Figure 1 shows the schematic process flow diagram of the processing of a laterite-nickel ore acid leaching slag provided according to example 1 of the disclosure; and
Figure 2 shows the diagram of a side blown furnace device provided according to example 1 of the disclosure.
Wherein, the above accompanying drawings include the following reference symbols:
1: Smelting-melting zone; 2: Reducing zone; 3: Electrothermal settling zone; 101: Feeding port; 102: Smelting zone flue gas outlet; 103: Dual-channel spray gun; 201: Three-channel spray gun; 202: Reducing zone flue gas outlet; 301: Electrode; 302: Electrothermal settling zone flue gas outlet; 303: Smelting slag outlet; 304: Molten iron outlet.
Detailed Description of the Embodiments
It should be noted that the examples and features in the examples in the application can be combined with each other without conflicting. The disclosure will be described in detail below with reference to the drawings and in combination with examples.
As analyzed by the background technology, there are problems of low recovery rate of iron and poor quality of molten iron in the existing technology of recovering iron ore concentrate from laterite-nickel ore acid leaching slag. In order to solve the technical problems, the present disclosure provides a processing method of a laterite-nickel ore acid leaching slag.
In a typical embodiment of the application, a processing method of a laterite-nickel ore acid leaching slag is provided, and the processing method includes: step S1, oxidation roasting a raw material comprising a laterite-nickel ore acid leaching slag and a flux in a roasting kiln to obtain a high-temperature calcine and a roasting flue gas; step S2, in an oxygen-enriched air, preliminary smelting a raw material comprising the high-temperature calcine and a first fuel in the smelting-melting zone of the side blown furnace to obtain a melt and a smelting zone flue gas; step S3, in a reducing atmosphere, reduction smelting a raw material comprising the melt, a reducing agent, and a second fuel in the reducing zone of the side blown furnace to obtain a reduced melt and a reducing zone flue gas; and step S4, subjecting the reduced melt to slag-iron separation in the electrothermal settling zone of the side blown furnac to obtain a smelting slag, molten iron, and an electrothermal settling zone flue gas.
The laterite-nickel ore acid leaching slag and a solvent are sent to the roasting kiln for oxidation roasting, so as to desulphurize and dehydrate (where the dust returned from the smelting system can be fed together into the roasting kiln after batching). The desulfurization rate of the oxidation roasting is greater than 95%, and the chemical reactions of the sulfur elements involved in the oxidation roasting process are as follows:
Fe 2(SO 4 )3 = Fe 203+3SO2+1.502
Cr2 (SO 4)3 = Cr 203+3SO2+1.502
A12(SO 4) 3 = A1 2 03+3SO2+1.502
4FeS+110 2 = 2Fe 2O3+8SO2
The high-temperature calcine sent to the smelting-melting zone of the side blown furnace quickly melts under the large amount of heat provided by the fuel and the oxygen-enriched air, forming a melt, which is transferred to the reducing zone of the side blown furnace, and in the reducing atmosphere, the metals in the melt are deeply reduced by a reducing agent to obtain a reduced melt including metal and slag. Subsequently, the reduced melt flows into the electrothermal settling zone to further settle and separate the metal iron in the slag, after which the smelting slag is discharged through the slag notch and water quenched to obtain harmless glass slag, the molten iron is discharged through the metal notch and the ingot casting is sold, the recovery rate of iron is above 95%, and the sulfur content of the cast iron is less than 0.03%. The above processing method is conducive to improving the recovery rate of iron and the product quality of molten iron. The waste heat of flue gas generated from the whole process can be fully recovered, thereby reducing the energy consumption and improving the system heat utilization rate, and the harmless glass slag can be used as a building material, and the above processing method is simple, has low process costs and high economic benefits.
Preferably, the high-temperature calcine obtained from the roasting kiln is sent to the side blown furnace by use of an insulated tank for smelting, in order to minimize the energy consumption as much as possible.
Preferably, the above high-temperature calcine produced by the roasting kiln can be cooled by cooling equipment before being sent to the side blown furnace (side blown molten bath smelting device) for smelting. In areas with abundant electricity, an electric furnace smelting device can also be used instead of the side blown molten bath smelting device.
In one embodiment of the application, in the above step S3, the reducing atmosphere is a strong reducing atmosphere, which comprises carbon monoxide and carbon dioxide, and the volume ratio of carbon monoxide to the strong reducing atmosphere is 60-90:100, such as 60:100, 70:100, 80:100, and 90:100.
In the above reducing atmosphere, it is beneficial to improve the reduction efficiency and effect of metals, thereby reducing the iron in the laterite-nickel ore acid leaching slag to elementary iron as much as possible.
Preferably, in the above step S3, the mass of the reducing agent is 15-25% by weight of the laterite-nickel ore acid leaching slag on a dry basis, the reducing agent is preferably selected from any one or more of pulverized coal, coke powder, charcoal active powder, and graphite, which is conduce to ensuring the sufficient contact between the reducing agent and the melt, and reducing the iron ions therein to elementary iron as much as possible.
In one embodiment of the application, in the above step S3, the temperature of the reduction smelting is 1450-15500C, and the time of the reduction smelting is preferably 0.5-2.5 hours; and the second fuel is preferably selected from any one or more of natural gas, coal gas, pulverized coal, heavy oil, pyrolysis oil, and biomass oil.
The above reduction smelting conditions are conducive to improving the reduction efficiency and degree of iron oxide in the laterite-nickel ore acid leaching slag, thereby improving the recovery rate of iron and the product quality.
Preferably, in the above step S1, the temperature of the oxidation roasting is 800-1200C, and the desulfurization rate of the oxidation roasting is preferably greater than 90%, the waste heat is preferably recovered by passing the roasting flue gas through a waste heat boiler, which is then subjected to dust collection and purification before reaching the standard for discharge. The above oxidation roasting conditions are conducive to improving the efficiency of desulfurization and moisture removal in the laterite-nickel ore acid leaching slag. The composition of the laterite-nickel ore acid leaching slag is mainly composed of iron oxide, and preferably the above laterite-nickel ore acid leaching slag is more suitably used in the above processing method of the application to obtain high recovery rate and high quality of iron.
The above laterite-nickel ore acid leaching slag is not the commonly existing laterite-nickel ore acid leaching slag in the industry, and the content of the main metal element iron thereof is between 35% and 60%.
Preferably, the mass ratio of the flux to the laterite-nickel ore acid leaching slag is 5-20:100, the flux is preferably limestone and/or quicklime, and the mass ratio of CaO to SiO 2 in the smelting slag is preferably controlled to be 0.5-1.5:1.
The flux itself has acidity or alkalinity, the flux and the impurities in the laterite-nickel ore acid leaching slag form a low melting point slag through acid-base interaction, only in such a way that the impurities in the laterite-nickel ore acid leaching slag can be separated from the iron metal. Preferably, the mass ratio mentioned above is conducive to separating the iron metal from the laterite-nickel ore acid leaching slag as much as possible. Preferably, the use of limestone and/or quicklime as a flux is beneficial for better regulating the slag type generated by the laterite-nickel ore acid leaching slag and melting point thereof, improving the melting efficiency of the laterite-nickel ore acid leaching slag, and is conducive to reducing the costs of the flux. Controlling the mass ratio of CaO to SiO 2 in smelting slag is conducive to reducing the viscosity and melting point of the slag.
In one embodiment of the application, in the above step S2, the temperature of the preliminary smelting is 1450-15500C, the time of the preliminary smelting is preferably 0.5-2.5 hours, and the waste heat is preferably recovered by passing the smelting zone flue gas through a waste heat boiler, which is then subjected to dust collection and purification before reaching the standard for discharge.
The above preliminary smelting is conducive to melting the solid in the laterite-nickel ore acid leaching slag as much as possible, thus laying the foundation for reduction smelting, and improving the selective reduction of iron at lower reduction smelting temperatures, thus making preparations for the higher grades of iron ore concentrate. Preferably, the waste heat is recovered by passing the smelting zone flue gas through a waste heat boiler, which is conducive to improving the energy utilization, reducing energy consumption costs, and reducing environmental hazards.
Preferably, in the step S2, the oxygen content in the oxygen-enriched air is 50-100%, and the jetting pressure of the oxygen-enriched air is preferably 0.2-0.6 MPa; and the first fuel is preferably selected from any one or more of natural gas, coal gas, pulverized coal, heavy oil, pyrolysis oil, and biomass oil, which will provide sufficient heat for the preliminary smelting, allowing the laterite-nickel ore acid leaching slag to quickly melt, which in turn is conducive to reducing the reduction temperature of the reduction smelting process and thereby reducing the energy consumption. Moreover, under the above-mentioned jetting pressure and oxygen-enriched conditions, the laterite-nickel ore acid leaching slag has better reduction smelting effect, lower energy consumption, and higher selective reduction degree of iron. In the actual operation process, the oxygen-enriched air can be obtained by combining compressed air and oxygen.
Preferably, the above fuels come from a wide range of sources, which can reduce costs. Of course, those skilled in the art can also choose other suitable fuels, which will not be elaborated here.
In one embodiment of the application, the waste heat is recovered by a waste heat boiler after the above reducing zone flue gas is combined with the electrothermal settling zone flue gas, which is then subjected to dust collection and purification to obtain the coal gas, wherein 50-100% of the coal gas is sent back to the roasting kiln as a heat source of the roasting kiln, and the steam is producible from the remaining coal gas via a coal gas boiler; and preferably, the smelting zone flue gas, the reducing zone flue gas, the electrothermal settling zone flue gas, and the roasting flue gas are passed through the waste heat boiler to recover the steam produced during the process, which is combined with the steam produced by the coal gas boiler for power generation.
The flue gas waste heat produced during the above processing method is fully recovered, with low energy consumption and high system heat utilization rate.
Preferably, the smoke dust obtained from the electric dust collection of the smelting zone flue gas, the reducing zone flue gas, the electrothermal settling zone flue gas, and the roasting flue gas is granulated, and which is then returned to the roasting kiln as the dust return for the batching system of the roasting kiln.
In one embodiment of the application, in the above step S4, the temperature for the slag-iron separation in the electrothermal settling zone is 1450-1600C.
The above temperatures are conducive to improving the separation efficiency and effect of molten iron and smelting slag.
The beneficial effects of the application will be described below in combination with specific examples.
Example 1
The laterite-nickel ore acid leaching slag was processed by using the side blown furnace as shown in Figure 2 and according to the process flow as shown in Figure 1, the composition and content of the laterite-nickel ore acid leaching slag were as follows: Components TFe FeO Fe 2O 3 Ni Co Cr SiO 2 TiO 2 Al 2O 3
Content 51.38 0.53 72.81 _{0.16 0.0024 2.86 7.97 0.12 4.26 Components CaO MgO Mn K20 Na2O P S C Ig Content 0.038 0.56 0.22 0.0054 0.023 0.010 2.01 0.034 8.56
500000 t/a of laterite-nickel ore leaching slag and about 40000 t/a of quicklime were sent into the roasting kiln, and oxidation roasted at 1000°C for desulfurization and dehydration to obtain the high-temperature calcine, which was sent to the smelting-melting zone 1 of the side blown furnace through the feeding port 101 for preliminary smelting, the smelting melting was carried out using a dual-channel spray gun 103, with oxygen-enriched air jetted into the inner channel of the spray gun and pulverized coal jetted into the outer channel, the temperature of the smelting-melting zone 1 was controlled at 1450°C to obtain the melt, and the smelting zone flue gas was obtained at the smelting zone flue gas outlet 102. The melt flowed to the reducing zone 2, which was equipped with a three-channel spray gun 201 and a reducing zone flue gas outlet 202. The inner channel of the spray gun was jetted with 100000 t/a of bituminous coal reducing agent, the middle channel was jetted with oxygen-enriched air (with an oxygen content of 80% and a jetting pressure of 0.5 MPa), and the outer channel was jetted with pulverized coal, the temperature of the reducing zone was controlled at 1450°C, and the reducing atmosphere Vco/V(co+co 2 ) = 80% to obtain the reduced melt and the reducing zone flue gas. The reduced melt flowed to the electrothermal settling zone 3, and was heated to 1500°C via the electrode 301 in the electrothermal settling zone, about 250000 t/a of molten iron was obtained at the molten iron outlet 304, while the smelting slag obtained at the smelting slag outlet 303 was harmless slag, the recovery rate of iron was about 97.3% and the sulfur content of molten iron was about 0.01%, and the electrothermal settling zone flue gas was obtained at the electrothermal settling zone flue gas outlet 302.
The waste heat was recovered by passing the smelting zone flue gas through the waste heat boiler, which was then subjected to dust collection and purification before reaching the standard for discharge. The waste heat was recovered by the waste heat boiler after the reducing zone flue gas was combined with the electrothermal settling zone flue gas, which was then subjected to dust collection and purification to obtain about 506000 t/a of the coal gas, wherein 65% of the coal gas was sent back to the roasting kiln as the heat source of the roasting kiln, and the remaining coal gas was sent to the coal gas boiler to produce steam. The smelting zone flue gas, the reducing zone flue gas, and the electrothermal settling zone flue gas were passed through the waste heat boiler to recover the steam produced during the process, which was combined with the steam produced by the coal gas boiler for power generation.
Example 2
The difference from example 1 lied in that the bituminous coal reducing agent was 90000 t/a, and about 242000 t/a of molten iron was finally obtained.
Example 3
The difference from example 1 lied in that the bituminous coal reducing agent was 125000 t/a, and about 252000 t/a of molten iron was finally obtained.
Example 4
The difference from example 1 lied in that the temperature of the oxidation roasting was 800 °C, the reducing atmosphere Vco/V(co+co 2)=90%, and about 251000 t/a of molten iron was finally obtained.
Example 5
The difference from example 1 lied in that the temperature of the oxidation roasting was 1200 °C, the reducing atmosphere Vco/V(co+co2 )=60%, and about 248000 t/a of molten iron was finally obtained.
Example 6
The difference from example 1 lied in that the temperature of the preliminary smelting was 1550 °C, the temperature of the reduction smelting was 1550 °C, it was heated to 1600 °C via the electrode 301 of the electrothermal settling zone, and about 253000 t/a of molten iron was finally obtained.
Example 7
The difference from example 1 lied in that 500000 t/a of the laterite-nickel ore leaching slag and about 30000 t/a of quicklime were sent into the roasting kiln, and oxidation roasted at 900 °C for desulfurization and dehydration. The obtained high-temperature calcine was sent to the side blown furnace, the temperatures of both the smelting-melting zone and the reducing zone were controlled at 1450 °C, the reducing zone VcO/V(co+co 2 )=80%, the injection amount of the bituminous coal reducing agent in the reducing zone was 95000 t/a, and the temperature in the electrothermal settling zone was 1500°C, about 246000 t/a of molten iron was obtained, and about 408000 t/a of purified coal gas was collected, of which 73% was used for roasting in the roasting kiln, and the rest of which was sent to the coal gas boiler for steam production.
Comparative example 1
The difference from example 1 lied in that the laterite-nickel ore acid leaching slag was not subjected to oxidation roasting, and it was directly subjected to preliminary smelting, reduction smelting and separation in the side blown furnace, the reducing zone was controlled at VcON(coco 2 )=50%, the injection amount of the bituminous coal reducing agent was 80000 t/a in the reducing zone, and about 235000 t/a of molten iron was finally obtained.
The sulfur content of molten iron and the recovery rate of iron obtained from the above example 1 to 7 and comparative example 1 are listed in Table 1.
Table 1 Examples/Comparative Sulfur content of Recovery rate of examples molten iron (%) iron (%) Example 1 0.01 97.3 Example 2 0.01 95.4 Example 3 0.01 98.1 Example 4 0.02 97.7 Example 5 0.01 96.5 Example 6 0.01 98.4 Example 7 0.01 95.7 Comparative example 1 0.9 91.5
As can be seen from the above description that the above examples of the disclosure achieve the following technical effects:
The laterite-nickel ore acid leaching slag and a solvent are sent to the roasting kiln for oxidation roasting, so as to desulphurize and dehydrate (where the dust returned from the smelting system can be fed together into the roasting kiln after batching). The desulfurization rate of the oxidation roasting is greater than 95%, and the chemical reactions of the sulfur elements involved in the oxidation roasting process are as follows:
Fe 2(SO 4 )3 = Fe 2O 3 +3SO2+1.502
Cr2 (SO 4)3 = Cr 203+3SO2+1.502
A12(SO 4) 3 = A1 2 03+3SO2+1.502
4FeS+110 2 = 2Fe 2O3+8SO2
The high-temperature calcine sent to the smelting-melting zone of the side blown furnace quickly melts under the large amount of heat provided by the fuel and the oxygen-enriched air, forming a melt, which is transferred to the reducing zone of the side blown furnace, and in the reducing atmosphere, the metals in the melt are deeply reduced by a reducing agent to obtain a reduced melt including metal and slag. Subsequently, the reduced melt flows into the electrothermal settling zone to further settle and separate the metal iron in the slag, after which the smelting slag is discharged through the slag notch and water quenched to obtain harmless glass slag, the molten iron is discharged through the metal notch and the ingot casting is sold, the recovery rate of iron is above 95%, and the sulfur content of the cast iron is less than 0.03%. The above processing method is conducive to improving the recovery rate of iron and the product quality of molten iron. The waste heat of flue gas generated from the whole process can be fully recovered, thereby reducing the energy consumption and improving the system heat utilization rate, and the harmless glass slag can be used as a building material, and the above processing method is simple, has low process costs and high economic benefits.
The above contents only describe the preferred examples of the disclosure, and are not intended to limit the disclosure. For those skilled in the art, various modifications and changes can be made to the disclosure. Any modifications, equivalent substitutions, improvements, and the like made within the spirit and principle of the disclosure shall be included within the scope of protection of the disclosure.
Claims (19)
1. A processing method of a laterite-nickel ore acid leaching slag, wherein the processing method comprises the steps of:
step S1, oxidation roasting a raw material comprising a laterite-nickel ore acid leaching slag and a flux in a roasting kiln to obtain a high-temperature calcine and a roasting flue gas;
step S2, in an oxygen-enriched air, preliminary smelting a raw material comprising the high-temperature calcine and a first fuel in the smelting-melting zone of the side blown furnace to obtain a melt and a smelting zone flue gas;
step S3, in a reducing atmosphere, reduction smelting a raw material comprising the melt, a reducing agent, and a second fuel in the reducing zone of the side blown furnace to obtain a reduced melt and a reducing zone flue gas; and
step S4, subjecting the reduced melt to slag-iron separation in the electrothermal settling zone of the side blown furnac to obtain a smelting slag, molten iron, and an electrothermal settling zone flue gas.
2. The processing method according to claim 1, wherein in the step S3, the reducing atmosphere is a strong reducing atmosphere, which comprises carbon monoxide and carbon dioxide, and the volume ratio of carbon monoxide to the strong reducing atmosphere is 60-90:100.
3. The processing method according to claim 2, wherein in the step S3, the mass of the reducing agent is 15-25% by weight of the laterite-nickel ore acid leaching slag on a dry basis.
4. The processing method according to claim 3, wherein the reducing agent is selected from any one or more of pulverized coal, coke powder, charcoal active powder, and graphite.
5. The processing method according to any one of claims 1 to 4, wherein in the step S3, the temperature of the reduction smelting is 1450-1550 0C, and the time of the reduction smelting is preferably 0.5-2.5 hours.
6. The processing method according to claim 3, wherein the second fuel is selected from any one or more of natural gas, coal gas, pulverized coal, heavy oil, pyrolysis oil, and biomass oil.
7. The processing method according to any one of claims 1 to 4, wherein in the step S1, the temperature of the oxidation roasting is 800-1200C.
8. The processing method according to claim 7, wherein the desulfurization rate of the oxidation roasting is greater than 90%, the waste heat is recovered by passing the roasting flue gas through a waste heat boiler, which is then subjected to dust collection and purification before reaching the standard for discharge.
9. The processing method according to any one of claims 1 to 4, wherein the mass ratio of the flux to the laterite-nickel ore acid leaching slag is 5-20:100.
10. The processing method according to claim 9, wherein the flux is limestone and/or quicklime, and the mass ratio of CaO to SiO 2 in the smelting slag is controlled to be 0.5-1.5:1.
11. The processing method according to any one of claims 1 to 4, wherein in the step S2, the temperature of the preliminary smelting is 1450-15500C, the time of the preliminary smelting is 0.5-2.5 hours.
12. The processing method according to claim 11, wherein the waste heat is recovered by passing the smelting zone flue gas through a waste heat boiler, which is then subjected to dust collection and purification before reaching the standard for discharge.
13. The processing method according to claim 11, wherein in the step S2, the oxygen content in the oxygen-enriched air is 50-100%.
14. The processing method according to claim 13, wherein the jetting pressure of the oxygen-enriched air is 0.2-0.6 MPa.
15. The processing method according to claim 14, wherein the first fuel is selected from any one or more of natural gas, coal gas, pulverized coal, heavy oil, pyrolysis oil, and biomass oil.
16. The processing method according to any one of claims 1 to 4, wherein the waste heat is recovered by a waste heat boiler after the reducing zone flue gas is combined with the electrothermal settling zone flue gas, which is then subjected to dust collection and purification to obtain the coal gas, wherein 50-100% of the coal gas is sent back to the roasting kiln as a heat source of the roasting kiln, and the steam is producible from the remaining coal gas via a coal gas boiler.
17. The processing method according to claim 16, wherein the smelting zone flue gas, the reducing zone flue gas, the electrothermal settling zone flue gas, and the roasting flue gas are passed through the waste heat boiler to recover the steam produced during the process, which is combined with the steam produced by the coal gas boiler for power generation.
18. The processing method according to claim 16, wherein the smoke dust obtained from the electric dust collection of the smelting zone flue gas, the reducing zone flue gas, the electrothermal settling zone flue gas, and the roasting flue gas is granulated, and which is then returned to the roasting kiln.
19. The processing method according to any one of claims 1 to 4, wherein in the step S4,
the temperature for the slag-iron separation in the electrothermal settling zone is 1450-1600 °C.
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| CN202211522059.8A CN116004936B (en) | 2022-11-30 | 2022-11-30 | Treatment method of laterite nickel ore acid leaching slag |
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| CN116949282B (en) * | 2023-04-28 | 2024-02-13 | 浙江华友钴业股份有限公司 | Method and equipment for treating laterite nickel ore leaching slag |
| CN116875759A (en) * | 2023-06-13 | 2023-10-13 | 中南大学 | A resource recovery method for recovering iron from high-pressure leaching residue of laterite nickel ore |
| WO2025059886A1 (en) * | 2023-09-20 | 2025-03-27 | 广东邦普循环科技有限公司 | Resource utilization method for iron-aluminum slag and nickel laterite ore acid leaching residue |
| CN117587226B (en) * | 2023-10-09 | 2025-02-14 | 浙江华友钴业股份有限公司 | Method and system for resource utilization of laterite nickel ore leaching tailings |
| CN117531821B (en) * | 2023-11-17 | 2024-10-11 | 宁波力勤资源科技股份有限公司 | Method and device for deep desulfurization of laterite wet smelting slag |
| CN117403057B (en) * | 2023-12-14 | 2024-03-08 | 中国恩菲工程技术有限公司 | Treatment method of laterite nickel ore acid leaching slag and active material |
| CN117721325B (en) * | 2024-02-07 | 2024-05-14 | 矿冶科技集团有限公司 | Method for extracting nickel-cobalt-iron from laterite-nickel ore |
| CN118222860B (en) * | 2024-02-28 | 2025-06-24 | 昆明理工大学 | Method for preparing nickel-iron alloy from laterite-nickel ore |
| CN117947261B (en) * | 2024-03-26 | 2024-05-28 | 扬州一川镍业有限公司 | Method for treating laterite-nickel ore leaching slag by using suspension magnetization roasting |
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| CN110331283B (en) * | 2019-08-19 | 2021-08-31 | 中国恩菲工程技术有限公司 | Treatment method of acid leaching residue of laterite nickel ore |
| CN114686699A (en) * | 2022-03-21 | 2022-07-01 | 贺毅林 | A kind of smelting process of laterite nickel ore |
| CN115386738B (en) * | 2022-08-10 | 2023-12-12 | 广东邦普循环科技有限公司 | Method for producing high nickel matte by reducing, vulcanizing and smelting laterite-nickel ore |
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