Hollow electrode metallurgical production process

Abstract

The present invention provides a metallurgical production process using a hollow electrode, which includes the following steps: 1. Calcining and drying the raw materials; 2. Pre-reducing and melting: feeding the calcined and dried raw materials into the hollow electrode, and blowing oxygen-rich gas and fuel from a combined oxygen lance located at the center of the electrode. The oxygen-rich gas reacts with carbon, hydrogen and metal oxides in the raw materials and fuel to release heat and generate reducing gas. The reducing gas contacts the raw materials countercurrently from bottom to top to reduce the oxides in the raw materials and generate metals. 3) Continuously adding materials. After the materials are melted, coal is concentrated, and an annular solid phase region is formed within the hollow electrode component. The reduced metals fall and melt into a liquid state. After the liquid metal or metal slag passes through the solid phase region, it precipitates and enters the metal layer at the bottom of the furnace. When the remaining metal oxides pass through the solid layer, they continue to react with carbon in the solid phase region to generate metals. In order to achieve the effects of short process flow; integrating calcination, pre-reduction, smelting reduction, and separation of metal and slag depletion.

Description

Technical Field

[0001] The present invention relates to the technical field of metallurgy, and specifically to a production process for hollow electrode metallurgy. Background Art

[0002] The primary smelting of ferrous metals and the smelting of non-ferrous metals have developed very mature technologies after a long period of development. However, their production has inherent deficiencies, namely their strong dependence on metallurgical coke. Due to the uneven distribution of coking coal resources, and the coking industry is the main source of environmental pollution. The production process of coking coal is long, with a large amount of flue gas emissions and many pollutant emissions. With the increasing attention to environmental pollution and the implementation of very strict environmental protection policies by the state, requiring ultra-low emissions of SO2, CO2, and NO in flue gas emissions, it will inevitably drive up the price of coking coal, thereby increasing the production cost of metals. Therefore, enterprises around the world are developing smelting methods that do not use metallurgical coking coal, or using electric furnaces for production. For example, the direct smelting reduction project of Baowu Company uses natural gas or coal to replace coking coal as fuel to produce hot metal. However, for general companies, the engineering technology requirements are high and the investment is large, making it impossible to learn and follow up.

[0003] Currently, enterprises are pursuing a production process with low equipment investment, low production cost, low operating expenses, simple and controllable production process, and low pollutant emissions. However, there is no existing production process on the market that can achieve the above goals. Summary of the Invention

[0004] In order to solve the technical problems in the background art, the present application provides a production process for super-large hollow electrodes, which adopts a super-large hollow electrode with an auxiliary combined oxygen lance to integrate pre-reduction and melting reduction, and can produce hot metal and non-ferrous metals.

[0005] The present patent application provides a production process for hollow electrode metallurgy, including the following steps:

[0006] 1) Calcining and drying the raw materials;

[0007] 2) Pre-reducing and melting;

[0008] Feeding the calcined and dried raw materials into the hollow electrode. The raw materials freely fall from the upper part of the hollow electrode under the action of their own weight. The oxygen-rich gas and fuel blown out by the combined oxygen lance located in the center of the electrode react with the carbon, hydrogen, and metal oxides in the raw materials and fuel to release heat and generate reducing gas. The released heat melts the raw materials, and the reducing gas contacts the raw materials countercurrently from bottom to top to reduce the oxides in the raw materials and generate metals.

[0009] 3) With the continuous addition of materials, after the materials melt, coal is enriched, forming an annular solid phase region inside the hollow electrode component. The reduced metal falls under its own weight and melts into a liquid state. After the liquid metal or metal slag passes through the solid phase region, it precipitates into the metal layer at the bottom of the furnace. When the remaining metal oxides pass through the solid layer, they continue to react with the carbon in the solid phase region to form metal.

[0010] In step 1, the raw materials and coal particles are calcined and dried. In step 2, the calcined and dried raw materials and coal particles are fed into the hollow electrode. The oxygen-rich gas blown out by the combined oxygen lance located at the center of the electrode reacts with the carbon, hydrogen, and metal oxides in the raw materials and coal particles to release heat and generate reduction gas.

[0011] In step 1, the raw materials, fluxes, and coal particles are calcined and dried. In step 2, the calcined and dried raw materials, fluxes, and coal particles are fed into the hollow electrode.

[0012] Step 1 specifically is: The raw materials are fed into the rotary kiln through a conveyor. The rotary kiln calcines the raw materials in a countercurrent manner. The required temperature of the discharged flue gas is ≤168°C. The water content of the calcined materials is ≤2%, and the remaining temperature of the calcined materials is 350°C ± 10°C.

[0013] The raw materials are transported and lifted by a belt conveyor and fed into the rotary kiln through a screw conveyor. The rotary kiln calcines the raw materials in a countercurrent manner. The rotational speed of the rotary kiln is 1 - 3 revolutions per minute. The materials stay in the rotary kiln for no less than 3 hours. The head of the kiln is high for feeding materials, and the tail is low for discharging materials. The required temperature of the flue gas discharged from the head of the kiln is ≤168°C, which is convenient for using reliable and economically priced filter bags. The water content of the calcined materials is ≤2%, and the remaining temperature of the calcined materials is 350°C ± 10°C.

[0014] The heat for calcination in step 1 comes from the physical heat in the pre-reduced flue gas in step 2 and the heat generated by secondary combustion. The CO and H2 contained in the flue gas are burned in the secondary fuel chamber added at the tail of the kiln before the flue gas enters the rotary kiln.

[0015] In step 2, the hollow electrode is inserted into the slag surface, and a positive pressure is formed inside the hollow electrode by using the slag liquid seal.

[0016] In step 2, the combined oxygen lance includes a central oxygen delivery pipe, an air pipe, a fuel pipe, and a forced water cooling pipe, or the combined oxygen lance consists of a central oxygen delivery pipe and an outer air cooling pipe body.

[0017] The combined oxygen lance is installed and aligned with the center point of the hollow electrode. The combined oxygen lance is insulated from the electrode during installation and use. The tip of the combined oxygen lance is 500 - 800 mm away from the slag surface.

[0018] A metallurgical electric furnace comprises a furnace wall and a furnace cover, wherein a hollow electrode is installed on the furnace cover, a combined oxygen lance is installed at the center point of the hollow electrode, the combined oxygen lance comprises a material pipe and a cooling pipe, the material pipe comprises a central oxygen supply pipe, an air pipe, and a fuel pipe, and the cooling pipe comprises a forced water cooling pipe or an outer air cooling pipe.

[0019] The benefits of this application are:

[0020] 1. The process has a short flow plug, integrating calcination, pre-reduction, smelting reduction, metal and slag depletion separation, and has the characteristics of concentrated energy, low energy consumption, large production capacity, low production cost and high efficiency.

[0021] 2. The production raw materials can be in block, granular or powder form, and the application range is very wide, and it can be applied to the melting of ferrous metals or non-ferrous metals.

[0022] 3. The fuel can be solid (coal-based), liquid or gaseous. The fuel and oxygen-rich air burn to generate high-temperature and high-heat molten materials and reduced oxides. Electric auxiliary heating is used to ensure metal reduction and separation. The operation is precise and simple. Electricity can be used for single reduction, metal melting and separation, ensuring safe production.

[0023] 4. Pre-reduction melting, molten metal oxidation and reduction gas can be controlled to ensure accurate and controllable reaction conditions.

[0024] 5. Oxygen-enriched blowing is used, with less flue gas volume and higher emission concentration. Especially in the smelting of non-ferrous metals, sulfur recovery equipment in the flue gas can be added to facilitate recovery, protect the environment and increase economic benefits.

[0025] 6. The rear end of pre-reduction and electric flue gas emission is equipped with a rotary dense calcination and drying of raw materials. The waste heat is fully utilized. The equipment is fully closed, with less air leakage, small flue gas volume, low production energy, economic and environmental protection.

[0026] 7. The equipment adopts conventional equipment, which is easy to maintain, requires less equipment investment, has a high degree of automation, and a good production environment. The production line operation industry is in line with the mainstream of modern industrial development.

[0027] 8. The whole process is fully sealed and mechanized. There is no irregular emission of waste gas and dust as in traditional industrial production. It complies with the requirements of modern environmental protection industry policies and is particularly suitable for the disposal of hazardous solid (liquid) waste. BRIEF DESCRIPTION OF THE DRAWINGS

[0028] Figure 1 It is a schematic diagram of the process of the present invention;

[0029] Figure 2 It is a structural schematic diagram of the electric furnace of the present invention;

[0030] 1 - Furnace wall, 2 - Furnace cover, 3 - Hollow electrode, 4 - Combined oxygen lance, 5 - Dilute phase zone, 6 - Dense phase zone, 7 - Solid phase zone, 8 - Bottom metal layer of the furnace, 9 - Slag layer. Specific implementation mode

[0031] The following will refer to the accompanying drawings to describe the preferred embodiments of the present application in detail. In order to enable those skilled in the art to better understand the solution of the present application, the technical solutions in the embodiments of the present application will be clearly and completely described below in conjunction with the drawings in the embodiments of the present application. The following embodiments are only used to illustrate the present invention and are not used to limit the scope of the present invention. For the experimental methods without specific conditions noted in the embodiments, they are usually carried out under conventional conditions or according to the conditions recommended by the manufacturer.

[0032] Example 1

[0033] The production process can be divided into three stages according to the material flow

[0034] 1. Material calcination and drying stage

[0035] The production raw powder, flux, and coal particles (powder) are put into the raw material bin after pretreatment. The production raw materials are required not to stick into lumps. When the moisture content is too high or the viscosity is too high, drying treatment is carried out. The flux can be selected to be added or not added according to the raw material analysis results. The composition should meet the requirements of the slag phase, and the general components are CaO, SiO2, etc. It can be added alone or in combination. The shape can be powder or granular. Add according to the process requirements. The fuel is coal particles, and coke fines can be used to replace coal particles under special requirements.

[0036] The bottom of the bin is distributed to the respective belt electronic scales by a disc feeder, and the feeding amounts of raw materials, flux, and fuel can be accurately controlled.

[0037] The feeding amount is determined according to the requirements of the production process. Raw materials, etc. are conveyed and lifted by belt conveyors and fed into the rotary kiln through screw conveyors. The rotary kiln uses countercurrent calcination of raw materials, which can make full use of heat energy. The rotational speed of the rotary kiln is 1 - 3 revolutions per minute. The residence time of the material in the rotary kiln is not less than 3 hours. The head of the kiln is high for feeding materials, and the tail of the kiln is low for discharging materials. The temperature of the flue gas discharged from the kiln head is required to be not greater than 168 °C, which is convenient for using reliable and economically priced filter bags. The moisture content of the material is not greater than 36%. This is mainly considered because some ore powders contain crystal water. Relaxing the moisture content of the material facilitates the diversity of raw material selection. The moisture content of the calcined material is not greater than 2%, and the remaining temperature is about 350 °C. It is fed into the electric furnace by a hoist. There is an intermediate heat-insulated high-level storage bin, and the capacity of the intermediate heat-insulated storage bin is about the amount used in 1 hour of normal production. The principle for the remaining temperature of the material is not to ignite the coal particles. Due to the large variability of the production capacity of the electric furnace and the combined production of normal electric furnace production capacity and the addition of auxiliary oxygen lances, the production capacity will be expanded by 1.5 times. Therefore, a margin should be left in the design of each equipment and the rotary kiln to ensure normal production. The flue gas flow rate in the rotary kiln is 0.5 - 1 m / s. The material filling rate is 20%, the filling angle is 120°, the rotational speed is 1.2 revolutions per minute, and the specific gravity of the material is 1.5 t / m 3 Calculate, and it can be adjusted according to the actual production situation if there are differences.

[0038] The kiln body rotates slowly. The raw material fine coal particles (powder) and flux powder are carried up by the kiln body under the action of friction. After exceeding the material movement angle, they flow down from the pile tip to the bottom feet under the action of gravity. Due to the inclined angle of the kiln body, the material will also move forward. The material tumbles continuously and stirs forward. The material is heated evenly, and the heat transfer resistance is small. If the raw material is in powder form and has a suitable moisture content, it will automatically form small balls, which is also one of the purposes of calcination. The production raw materials can be selected according to the actual situation and the production process can choose the state of pulverized coal and flux. When necessary, the raw material varieties also need to be adjusted accordingly to minimize the addition or not add flux as much as possible, select suitable raw materials containing SiO2, CaO, and FeO, reduce unnecessary furnace charge, thereby increasing the output and improving economic benefits.

[0039] The shape of the calcined product is preferably 5 - 25 mm. The moisture content shall not be higher than 2% and the powder content shall not be greater than 15%. The bulk density is 1.5 t / m 3 。

[0040] The heat for calcination comes from the physical heat in the furnace flue gas and the pre-reduced flue gas. Since CO and H2 are contained in the flue gas, before the flue gas enters the rotary kiln, it is burned in the secondary fuel chamber added at the kiln tail. The flue gas is mixed with the added fresh air and burned, and the supplied air enables CO and H2 to burn fully and generate heat, reheating the flue gas again. The high-temperature flue gas contacts the material, raising the temperature of the material for calcination and drying the material. At the same time, some parts of the raw materials are not thermally decomposed, and the volatiles react with C and the Fe2O3 contained in the material to form Fe3O4. Considering that too high a temperature will cause the coal to burn, the material temperature is increased on the premise of not causing the coal to burn. Since the flue gas temperature at the tail of the rotary kiln is much higher than that of the material, and the reduction reaction conditions of Fe2O3 are low and the reaction rate is fast, Fe3O4 can be formed at a temperature below 500 °C, and the reduction reaction will be very thorough. Due to the high moisture content of the raw materials, the moisture cannot be completely removed only by drying, so the calcination plus drying process is adopted.

[0041] The main harmful substances in the flue gas are CO2 and SO2. When it is necessary to recover them in the presence of alkali metals or zinc vapor, a cooling and cyclone dust collector can be added in front of the secondary combustion chamber to recover the metals. When the SO2 concentration is high and the gas volume is stable in non-ferrous metal smelting and needs to be recycled, sulfur recovery equipment such as sulfur-making equipment can also be installed in the flue gas in front of the secondary combustion chamber. After the flue gas enters the secondary combustion chamber, it burns fully in the excess air, re-oxidizing the metal oxides and other harmful substances in the flue gas, making the oxidation of the materials in the flue gas more complete, which is particularly important during hazardous waste disposal. Sometimes, due to insufficient oxidation during hazardous waste disposal, the discharged flue gas fails to meet the standards, causing secondary pollution.

[0042] Due to the adoption of oxygen-enriched blowing, the equipment has good sealing performance and less flue gas discharge. Compared with open electric furnaces and blast furnaces, the flue gas discharge volume is 50 - 60% less than that of the same production capacity of the same specification. Therefore, the investment in flue gas treatment equipment is small and the treatment is convenient.

[0043] After the flue gas is discharged from the rotary kiln, it enters the bag filter. After dust removal treatment, it enters the fan for pressurization, and then passes through the desulfurization and dust removal tower for dust removal, desulfurization, demisting and dehydration, and is discharged after meeting the standards.

[0044] 2. Pre-reduction and melting

[0045] Pre-reduction and melting are the second stage of the material reaction. The raw materials are quantitatively fed into the hollow electrode 3 through the screw conveyor at the bottom of the intermediate silo according to the production process requirements, and the raw materials fall freely due to their own weight. At this time, the electrode is used as the reaction reduction container shell, and the auxiliary combined oxygen lance 4 is inserted into the center of the hollow electrode 3 from above. The gas and fuel blown out by the combined oxygen lance 4 react with the carbon, hydrogen and metals in the slag to generate high temperature and reducing gas. The high-temperature heat melts the raw materials. The high-temperature reducing gas contacts the raw materials countercurrently from bottom to top, reducing the oxides in the raw materials to form metals and raising the temperature of the raw materials.

[0046] 2.1 Combined Electrode

[0047] The hollow electrode 3 has two functions: 1. Like conventional electrodes, it generates heat by heating with electricity. 2. The hollow electrode 3 is used as a container for pre-reaction reduction. Since the hollow electrode 3 is inserted about 500mm into the slag surface, the hollow electrode 3 is sealed with slag liquid, so that the inside and outside of the hollow electrode 3 can withstand a pressure difference of 15kPa, so that the high-temperature reaction inside the hollow electrode 3 does not affect the outside. At the same time, the high-temperature flue gas flows in the order of process requirements to heat the material. And the oxides in the material are reduced to metals. Due to the intense reaction of oxygen and carbon inside the hollow electrode 3, the local temperature generated is very high, and the heating temperature of the hollow electrode 3 is superimposed. It can accelerate the melting of raw materials. The high-temperature FeO slag after melting forms a strong scouring on the furnace wall. Reduce the life of the furnace wall 1. In actual production, it is found that the life of the furnace wall 1 is very low. Seriously affect normal production. In this patent application, the electrode is used as a container. It can withstand high temperature and heat, and is resistant to mechanical scouring of slag. And the electrode itself is a consumable, and the loss is 1 times greater than normal. Due to the use of self-baking electrodes, the price is low, the cost performance is high, and it is economical and practical.

[0048] Hollow electrode 3 When the hollow electrode 3 is not heated by the combined oxygen gun 4, the working voltage is increased, and the hollow electrode 3 is used in the same way as other solid electrodes. It is just that the electrode is hollow for charging and melting, which is more suitable for powder production. If the atmosphere in the central melting zone is an oxidizing atmosphere, an anti-oxidation layer can be applied on the inner surface of the electrode to protect the electrode and slow down the oxidation rate.

[0049] 2.2 Combined oxygen gun

[0050] The combined oxygen lance 4 mainly consists of a central oxygen delivery pipe, an air pipe, a fuel pipe and a forced water cooling pipe body, and can also consist of a central oxygen delivery pipe and an outer air cooling pipe body. Suitable products are selected according to process requirements.

[0051] The combined oxygen lance 4 must be installed to align with the center of the electrode and ensure that the oxygen lance does not deviate much during use. The combined oxygen lance 4 must be insulated and safe during installation and use. The end of the combined oxygen lance 4 is about 500-800mm away from the slag surface to ensure safety. The heat generated by the oxygen blowing reaction of the combined oxygen lance 4 is required to be just balanced with the heat of material melting. Most of the coal-based fuel used for the reaction is added before the calcination rotary kiln. The insufficient part is added with the assistance of the oxygen lance. Some materials cannot be added with fuel before the rotary kiln, so they are sprayed in through the combined oxygen lance 4 pipeline.

[0052] The coal powder used for injection of the combined oxygen lance 4 needs to be dried to a moisture content of no more than 1%, ground into a size of less than 200 meshes and sent to the coal bunker for standby use.

[0053] When using liquid fuel, clean the pipe by blowing it with dry air before use.

[0054] In special cases, the combined oxygen lance 4 can be lined with refractory materials on the outer pipe sleeve to protect the oxygen lance. The flow rate of pipeline oxygen should not be greater than 60 m / s, and the flow rate of the oxygen lance nozzle is 150 m / s. One oxygen lance is in use and one is in reserve to ensure timely replacement in case of a fault with the oxygen lance.

[0055] 2.2.3 Material state process

[0056] The combined oxygen lance 4 is inserted from the center of the hollow electrode 3 to a distance of 500 - 800 mm from the slag surface, and air and oxygen are fed in. The air and oxygen are mixed into oxygen-enriched gas, which reacts with the coal in the raw materials to generate high temperature and heat, melting the materials, and forming a cavity and a small molten pool. The strongly reactive high-temperature gas generates a certain positive pressure, and the high-speed blown gas causes the molten slag to circulate at high speed in the molten pool, accelerating the melting speed of the materials. Due to the continuous addition of materials and the enrichment of coal after the materials are melted, an annular solid phase zone 7 will be formed between the central molten pool and the electrode. The liquid of the high-temperature melted materials flows in balance with the outside slag liquid 9 through the voids in the solid phase zone 7. The formation of the solid phase zone 7 is beneficial to protecting the electrode consumption and the positive pressure in the high-temperature zone. Since the electrode is inserted into the slag surface to a depth of not less than 400 mm, a seal is formed. The slightly positive pressure is not greater than 15 KPa. The high-temperature and high-heat gas can only flow upward in the hollow electrode 3. In the production with small-particle raw materials, the materials are pre-reduced in the fluidized bed state, and the temperature gradient generated by the high-temperature reduction gas in the materials is not large, and the reduction speed of the materials is fast. When using lump raw materials for production, the materials work in the moving bed state. In the moving bed pre-reduction, the temperature gradient of the materials is large, the operation control is relatively easy, the tail gas emission temperature is low, and the energy consumption is small. Its working principle is similar to that of shaft furnace sintering, different from other direct smelting reduction technologies. Since the diameter of the hollow electrode 3 cannot be made very large, and the length (width) is 4 - 5 m. When the materials pass through the electrode, the pre-reduction can achieve a metallization rate of about 70% of the raw materials. The reduction time of the materials at a temperature above 850 °C should not be less than 10 minutes. Due to the small cross-section of the electrode, the production volume should be as large as possible. The gas generated by excessive chemical reactions will block the normal passage of the materials. To balance this relationship, only a certain amount of fossil fuel can be added to the raw materials to ensure that the heat energy and reduction gas volume required for pre-reduction reach a balanced relationship, and the other lost heat energy is supplied by electric heating.

[0057] Above the dense phase zone 6 or the moving bed, there is a dilute phase zone 5 composed of dust. The dilute phase zone 5 is determined by the dust particles and the flue gas flow rate. When the dust particle size is small and the flue gas flow rate is too high, the dust is carried away by the flue gas and lost. Therefore, the flue gas flow rate should not be too high to avoid material loss, and it is required that the flue gas flow rate under standard conditions is not greater than 4 m / s. The porosity of the materials in the dense phase zone 6 or the moving bed is 30%. The weight of the material layer in the dense phase zone 6 is equal to the product of the bed pressure and the cross-sectional area of the bed.

[0058] The raw materials calcined in the rotary kiln enter the interior of the electrode through a metering screw feeder in the intermediate bin, pass through the dilute phase zone 5 to the dense phase zone 6 or the moving bed, and are heated, reacted, and reduced of metal oxides through high-temperature reduction and heat exchange of the body. Due to the self-weight of the material, it slowly falls into the high-temperature melting zone and melts into a liquid state. In fact, the high-temperature gas flow generated by the carbon-oxygen reaction in the high-temperature zone at the center of the hollow electrode 3 rises rapidly and will carry part of the material upward, while the wind speed in the inner wall area of the hollow electrode 3 is relatively low, and the material sinks along the inner wall of the electrode, forming a circulation flow. Part of the material settles into the solid phase zone 7 to supplement the consumption of melting, and the other part is carried into the circulation by the gas flow and is reduced again. At the same time, since the material sinks along the inner wall of the electrode, a material layer is formed to protect the electrode. Since new materials are continuously added, the temperature of the material layer will not be too high. The circulating material transfers heat energy to the hollow electrode 3 to assist in the sintering of the hollow electrode 3.

[0059] Using a moving bed for pre-reducing materials can form a gradient with an obvious temperature difference, which is more suitable for this process, helps reduce production energy consumption, ensures normal production. The combined oxygen lance design can add pulverized coal for gasification of the combustion-supporting gas, can adjust the reduction atmosphere, and increases the applicable range of raw materials.

[0060] 2.3 Separation and precipitation of metal oxides

[0061] The high-temperature and overheated liquid metal slag passes through the solid layer, and the metal precipitates into the bottom metal layer 8 (matte layer). The remaining metal oxides in the slag continue to react with the carbon in the solid layer to form metal when passing through the solid layer. Due to the pressure difference, the liquid slag flows into the outer slag layer 9. Due to the special structure of the hollow electrode, the reaction on the outer layer of the electrode is relatively calm. The voltage and current used for the hollow electrode 3 are set, controlled, and adjusted according to the temperature of the slag and the temperature of the bottom metal layer 8 (matte layer) to ensure that the temperatures of the bottom metal layer 8 (matte layer) and the slag meet the requirements for normal production.

[0062] During normal production, the electrode voltage is relatively low, which is the same as the principle of the depleting furnace. The electricity consumption basically maintains the balance of the furnace consumption, and the electricity consumption is 150 - 200 Kwh / t. When the molten slag flows from the inner hollow layer to the outer layer of the hollow electrode 3, due to the gas contained in the molten slag not being discharged in time, the molten slag forms a foamy slag around the electrode, effectively preventing heat dissipation and protecting the furnace cover 2 and refractory materials.

[0063] According to the process requirements, part of the coke breeze and flux can be added to the molten slag. The metal oxides in the slag continue to be reduced with the added carbon to form metal, which grows and precipitates. Due to the reasonable space, time, temperature, and reduction and separation conditions provided by the outer layer of the electrode, the metal recovery in the slag is significantly higher than that in other electric furnaces or blast furnaces, minimizing the recoverable metal in the discharged waste slag.

[0064] The outer environment of the hollow electrode 3 is relatively calm, and the scouring of the slag on the refractory furnace wall will be greatly improved, which is conducive to significantly increasing the service life of the refractory furnace. The hollow electrode 3 operates in a high-current and low-voltage mode. In the area near the furnace wall, a solid slag surface will form on the slag surface, shortening the slag heat flow distance and intensity, effectively reducing heat dissipation and energy consumption. The slag surface is at normal pressure and low negative pressure, so normal-flow slag discharge can be adopted safely and conveniently. Metal discharge can be carried out through another opening in the conventional way.

[0065] III. Metallurgical Reactions and Heat Calculations

[0066] The treatment objects in the metallurgical industry are mainly iron oxides and iron compounds. Therefore, we conduct metallurgical analysis based on iron oxides. Since metallurgical reactions are very complex, we simplify the analysis and calculation to only the basic data required for normal production processes to maintain the heat balance required for the reaction and facilitate equipment selection.

[0067] To ensure the normal progress of the molten separation of metals and the reduction reaction of oxides, the production process must meet two necessary conditions: 1. The heat required for material melting; 2. The reducing agent dosage required for the metal reduction reaction and the reaction temperature. To visually analyze the metallurgical reaction situation of the materials, we take the recycling and reuse of the post-smelting dust in a certain factory as a case study.

[0068] Table 1 Analysis of the Composition of 100 kg of Materials

[0069] Substance NiO CuO Fe3O4 FeS MgO SiO2 CaO Others Total Mass kg 8.58 4.15 54.35 3.67 5.35 17.30 1.04 5.61 100 Molecular weight 74.7 79.54 231.55 87.92 40.30 60.1 56.1

[0070] Table 2 Relative Heat Capacities of Relevant Substances (KJ / (Kg.mol))

[0071] Substance NiO CuO Fe3O4 FeS MgO SiO2 CaO Relative heat capacity 80005 84904 290152 127100 71110 97811 75295

[0072] Table 3 Chemical Composition of Typical Reduction Coal (%) M is moisture, A is ash, V is volatile matter, and FC is the carbon content after volatilization

[0073] Chemical composition C H N O M A V Std FC Content 80.90 4.2 0.90 4.50 8.3 9.3 18.8 0.23 71.9

[0074] 3.1 Raw Material Analysis

[0075] Based on the data provided in Table 1 and Table 2, calculate the heat enthalpy absorbed by 100 Kg of raw materials with 0% moisture when the temperature rises from 298 K to 1723 K.

[0076] Table 4

[0077]

[0078] The average heat enthalpy of the raw materials is 133415.19 KJ / (1723K - 298K) = 0.936 KJ / Kg.K.

[0079] 3.1.1 The mass of the raw materials is 1000 kg, the temperature of the raw materials entering the furnace is 300 °C (573 K), the average melting temperature is 1450 °C (1723 K), the moisture content is 0%, and losses are not considered.

[0080] The heat energy required for the material is H = 1000 Kg * 0.936 * (1723 - 573 K) = 1076630 KJ

[0081] 3.2 Analysis of the reduction reaction of metal oxides in the raw materials

[0082] The main metal oxides in the principle are NiO, CuO, Fe3O4 and sulfide FeS. The main reaction equations are as follows:

[0083] FeS + 1.5O2 = FeO + SO2 (3 - A)

[0084] Fe3O4 + CO = 3FeO + CO2 (3 - B)

[0085] FeO + CO = Fe + CO2 (3 - C)

[0086] NiO + Fe = Ni + FeO (3 - D)

[0087] CuO + Fe = Cu + FeO (3 - E)

[0088] C + 2O2 = 2CO2 (3 - F)

[0089] 3.2.1 FeS reaction

[0090] Analyzed by the reaction equation (3 - A), the raw material is 1000 Kg, and FeS is 36.7 Kg (0.4174 Kmol)

[0091] FeS + 1.5O2 = FeO + SO2

[0092] ΔH f.298 = -468441 KJ / Kmol (FeS)

[0093] The heat of formation Q of the FeS reaction = (36.7 / 87.92) * (-468441) = -195539 KJ

[0094] The amount of FeO formed by the reaction is: (36.7 / 87.92) * 71.82 = 29.98 Kg

[0095] The amount of SO2 formed by the reaction is: (36.7 / 87.92) * 64.1 = 26.76 Kg

[0096] The demand for O2 is: (36.7 / 87.92) * 1.5 * 32 = 20.04 Kg

[0097] 3.2.2 Fe3O4 Reaction

[0098] The raw material is 1000 Kg, and Fe3O4 is 543.5 Kg (2.347 Kmol)

[0099] Fe3O4 + CO = 3FeO + CO2

[0100] ΔH f.298 = 19287 KJ / Kmol

[0101] The heat of formation of Fe3O4 reaction Q = 2.347 * 19287 = 45266.59 KJ

[0102] The amount of FeO formed in the reaction is: 2.347 * 3 * 71.82 = 505.9 Kg

[0103] The amount of CO2 formed in the reaction is: 2.347 * 44 = 103.27 Kg

[0104] The demand for CO is: 2.347 * 28 = 65.72 Kg

[0105] 3.2.3 NiO Reaction

[0106] The raw material is 1000 Kg, and NiO is 85.8 Kg (1.149 Kmol)

[0107] NiO + Fe = Ni + FeO

[0108] ΔH f.298 = -31464 KJ / Kmol

[0109] The heat of formation of NiO reaction Q = 1.149 * (-31464) = -36139.6 KJ

[0110] The amount of FeO formed in the reaction is: 1.149 * 71.85 = 82.53 Kg

[0111] 3.2.4 CuO Reaction

[0112] The raw material is 1000 Kg, and CuO is 41.5 Kg (0.522 Kmol)

[0113] CuO + Fe = Cu + FeO

[0114] ΔH f.298 = -116190 KJ / Kmol

[0115] The heat of formation of CuO reaction Q = 0.522 * (-116190) = -60622.1 KJ

[0116] The amount of FeO generated by the reaction is: 0.522 * 71.85 = 37.49 Kg

[0117] 3.2.5 FeO Reaction

[0118] The raw material is 1000 Kg. According to 3.2.1, 3.2.2, 3.2.3, and 3.2.4, the calculated FeO is 665.9 Kg (9.129 Kmol).

[0119] During metal smelting, the basicity of the slag is very important. According to the analysis in Table 1, no flux needs to be added to the raw materials of this application, and a part of FeO will melt into the slag and combine with SiO2. The slag basicity is 1.15.

[0120] In 1000 Kg, Mg is 5.35 Kg, SiO2 is 17.3 Kg, and CaO is 1.04 Kg.

[0121] The content of FeO in the slag is: 173 * 1.15 - 10.4 - 53.5 = 135.1 Kg

[0122] The actual amount of FeO generated to form metal is: 655.9 - 135.1 = 520.8 Kg (7.25 Kmol)

[0123] Calculate the consumption of the reducing agent using CO as the reducing agent.

[0124] FeO + CO = Fe + CO2

[0125] The amount of the reducing agent is 7.25 * 28 = 203 Kg

[0126] 3.2.6 Reducing Agent

[0127] When reducing Fe in 1000 Kg of raw materials to metallic iron with CO as the reducing agent.

[0128] The amount of the reducing agent is: 2.347 + 7.25 = 9.597 Kmol.

[0129] 3.3 Reducing Agent Analysis

[0130] Use coal as the fuel reducing agent. Since the coal has been calcined and dried in a rotary kiln before being fed into the furnace, the moisture content is counted as 0%. According to Table 3, C = 80.9%, H = 4.2%, N = 0.9%, O = 4.5%, A = 9.3%, Std = 0.23%. Analyzing 1 Kg of dry coal, the utilization rate of CO is set as Y CO = 6%, and the utilization rate of hydrogen Y H = 10.7%. The heat capacity of the slag is 1.75 KJ / Kg.K.

[0131] The calorific value of coal

[0132] Q = (1 - 9.3%)[(9203 + 41706 * 6%) * 80.9 + 132953 * 4.2% * 10.7%] = 9130.88 KJ / Kg。

[0133] Carbon content in coal: 0.809 / 12 * (1 - 9.3%) = 0.061 Kmol

[0134] Amount of CO: 0.06 * (1 - 6%) = 0.057 Kmol

[0135] Amount of CO2: 0.061 - 0.057 = 0.004 Kmol

[0136] Amount of H: 0.042 / 2 * (1 - 9.3%) = 0.019 Kmol

[0137] Amount of H2: 0.019 * (1 - 10.7%) = 0.017 Kmol

[0138] Amount of H2O: 0.019 - 0.017 = 0.002 Kmol

[0139] Amount of N2: 0.009 / 28 * (1 - 9.3%) = 0.00029 Kmol

[0140] Total reducing gas: 0.057 + 0.004 + 0.017 + 0.002 + 0.00029 = 0.0805 Kmol。

[0141] Table 5 Composition of reducing gas

[0142] Component CO <![CDATA[CO2]]> <![CDATA[H2]]> <![CDATA[H2O]]> <![CDATA[N2]]> Total Molar quantity 0.057 0.004 0.017 0.002 0.00029 0.0805 Mole percentage % 71.41 4.56 21.23 2.53 0.37 100

[0143] Physical heat loss of coal ash slag is: 0.093 * 1.75 * (1723 - 298) = 232 KJ / Kg;

[0144] Gas oxidation consumption: (0.061 + 0.004 + 0.002) / 2 = 0.0334 Kmol。

[0145] 3.3.1 Coal demand for reduction

[0146] For 1000 Kg of raw materials, the theoretical amount of reducing agent used is 9.597 Kmol. The amount of CO generated per kilogram of coal for reduction is 0.057 Kmol, and the amount of H2 is 0.017 Kmol. The main components participating in the reduction reaction are CO and H2.

[0147] Coal consumption: 9.597 / (0.057 + 0.017) = 129 Kg

[0148] The calorific value of 129 Kg of coal is 1177883.52 KJ

[0149] The amount of CO generated from 129 Kg of coal is 7.415 Kmol

[0150] The amount of CO2 generated from 129 Kg of coal is 0.473 Kmol

[0151] The amount of H2 generated from 129 Kg of coal is 2.194 Kmol

[0152] The amount of H2O generated from 129 Kg of coal is 0.263 Kmol

[0153] The amount of N2 generated from 129 Kg of coal is 0.0.38 Kmol

[0154] The amount of oxygen required for the combustion of 129 Kg of coal is 4.312 Kmol

[0155] The physical heat loss of the 129 Kg coal return slag is 29928 KJ

[0156] 3.3.1.1 Oxygen-enriched combustion support analysis

[0157] Oxygen-enriched gas is used to support the combustion of coal. A combination of pressurized air and pure oxygen is used, and the oxygen content in the mixed gas can be adjusted in the range of 30 - 60% to meet the requirements of various production processes.

[0158] Calculated based on an oxygen content of 60%:

[0159] The standard-state volume of oxygen required is: 4.938 Kmol * 22.4 m 3 / Kmol = 110.6 m 3

[0160] The standard-state volume of the mixed gas is: 110.6 / 60% = 184 m 3

[0161] The nitrogen content of the mixed gas is: 184 - 110.6 = 73.7 m 3

[0162] 73.7 m 3 / 22.24 m 3 / Kmol = 3.314 Kmol

[0163] The amount of N carried by 129 Kg is 0.038 Kmol, and the total nitrogen content is 0.038 + 3.314 = 3.352 Kmol.

[0164] The amount of air added in the combustion-supporting agent is 73.7 / 79% = 93.3 m 3

[0165] The amount of pure oxygen added in the combustion-supporting agent is 184 - 93.3 = 91 m 3 。

[0166] Table 6 Composition of the reducing gas produced after the combustion gasification reaction of 129 Kg of coal

[0167] Component CO <![CDATA[CO2]]> <![CDATA[H2]]> <![CDATA[H2O]]> <![CDATA[N2]]> Total Molar quantity 7.415 0.473 2.194 0.263 3.352 13.697 Mole percentage % 54.14 3.45 16.02 1.92 24.47 100

[0168] 3.3.2 Analysis of the reduction reaction of the reducing gas and the oxide

[0169] The amount of raw materials fed into the furnace is 1000 Kg, and the reducing gas in Table 7 is used to reduce the iron oxide. The steps of the smelting reduction of iron are: Fe2O3 - Fe3O4 - FeO - Fe. For Fe3O4, 2.347 Kmol of reducing gas is required, which is 7.25 Kmol in terms of the intermediate product FeO, calculated based on a conversion rate of 70% for FeO - Fe. The higher the conversion rate, the higher the corresponding requirements, which is not conducive to control. In the reduction reaction, H2 has a fast reaction rate and significantly higher diffusivity than CO. The thermodynamic comprehensive utilization rate of H2 increases with the increase in temperature. Assuming H2 is completely reacted and utilized, and the rest is reduced by CO gas.

[0170] Reduction reaction of FeO

[0171] The amount of FeO to be reduced is: 7.25 * 70% = 5.075 Kmol.

[0172] The amount of H2 involved in the reduction reaction is: 2.194 Kmol.

[0173] The amount of CO involved in the reduction reaction is: 5.075 - 2.194 = 2.881 Kmol.

[0174] The amount of H2O generated by the reaction is: 2.194 Kmol.

[0175] The heat absorbed by the reaction of FeO with H2 is: 2.194 * 30230 = 66329.49 KJ.

[0176] The amount of CO2 generated by the reaction of FeO with CO is: 2.881 Kmol.

[0177] The heat generated by the reaction of FeO with CO is: 2.881 * (-10920) = -31458.76 KJ.

[0178] The heat loss in the FeO reduction reaction is: 66329.49 + (-31458.76) = 34870.73 KJ.

[0179] Reduction reaction of Fe3O4

[0180] Calculated based on Fe3O4 at 2.347 Kmol, the amount of CO required for reduction is 2.347 Kmol, and the amount of CO2 generated is 2.347 Kmol.

[0181] The amount of CO in the tail gas after the Fe3O4 reduction reaction is: 7.415 - 2.881 - 2.347 = 2.187 Kmol.

[0182] The amount of CO in the tail gas after the reaction is: 0.473 + 2.881 + 2.347 = 5.7 Kmol.

[0183] The amount of H2O in the tail gas after the reaction is: 2194 + 0.263 = 2.457 Kmol.

[0184] Table 7 Composition of the tail gas after the reduction reaction

[0185] Component CO <![CDATA[CO2]]> SO2 <![CDATA[H2O]]> N2 Total Molar quantity 2.187 5.7 0.418 2.457 3.352 14.114 Mole percentage 15.5 40.39 2.96 17.40 23.74 100 Marked state volume 47.2 126.9 9.4 76.4 259.9 Mass 61.23 250.85 26.76 44.23 93.9 476.97 Average specific heat capacity 31.38 49.53 50.88 37.62 31.05 Unit heat loss 68.63 282.32 21.27 92.43 104.08 568.73

[0186] During normal production, the gas will automatically generate a balanced atmosphere. According to the process requirements, the tail gas emission temperature is 650 °C, the equilibrium coefficient K = 2.3, and it is speculated that CO = 1.387 Kmol, CO2 = 6.5 Kmol, H2 = 0.8 Kmol, and H2O = 1.657 Kmol. Therefore, attention should be paid to the tail gas detection in actual production.

[0187] Calculating and citing the data do not affect the subsequent calculations. Still cite the original data.

[0188] C: Tail gas

[0189] (1) Coal is burned and gasified with an oxygen-rich mixed gas to generate a reducing gas. The metal oxides in the raw materials are reduced through the raw material layer to generate high-temperature tail gas. The average temperature of the tail gas is 650 °C (923 K), the mass of the tail gas is 476.97 Kg, the standard volume is 260 m 3 (i.e., at 25 °C), and the standard specific gravity is 1.83 kg / m 3 . The average heat capacity is 568.73 KJ.K.

[0190] The physical heat carried by the tail gas is Q = 568.73×(923 - 298 k) = 355456.25 KJ.

[0191] (2) During the production of the electric furnace, due to production needs and equipment leakage, high-temperature flue gas will also be generated. Although the flue gas is not mixed with the above formula, it is mixed in the combustion chamber of the calcining rotary kiln when the flue gas enters. According to actual experience, the heat loss in the electric furnace flue gas is the same as 20% of the heat loss in the above formula flue gas.

[0192] The physical heat loss carried by the electric furnace flue gas is Q = 355456.25×20% = 71091.25 KJ

[0193] The standard volume of the flue gas is 260×20% = 52 m 3 [Total flue gas = 260 + 52 = 312 m 3

[0194] ​(3) The metallization rate of the process is set at 70%. That is, 70% of Fe is generated from the reduction reaction of FeO with CO. 30% of FeO in the remaining raw materials enters the slag after high-temperature melting. The carbon in the molten slag for the reduction reaction is partially from the remaining unreacted carbon in the raw materials and partially from the coke breeze added from the furnace cover after slag analysis, that is, the carbon added outside the electrode. The reaction is as follows:

[0195] FeO + C = Fe + CO H f298 = 161503 KJ / kmoL (FeO)

[0196] The amount of unreacted remaining FeO is 7.25×30% = 2.175 Kmol

[0197] The heat consumed by the reaction is 2.175×161503 = 351269 KJ

[0198] Note: The heat consumed in the above formula is not included in the carbon thermal balance calculation because it is not determined to be compensated by electricity.

[0199] The amount of CO gas generated by the reaction is 2.175 Kmol, and the total amount of CO in the tail gas is:

[0200] 2.187 + 2.175 = 4.362 Kmol

[0201] That is, the remaining gas generated by this reaction forms foamy slag around the electrodes, effectively protecting the high-temperature dissipation. The generated high-temperature furnace gas is discharged through the furnace cover and dust removal holes to the rotary kiln combustion chamber to recover the chemical energy in the flue gas.

[0202] 3.4 Comprehensive Thermal Balance Analysis

[0203] (1) Given that the weight of the raw materials entering the furnace is 1000 Kg, the temperature is 300 °C, the moisture content is 0%, no flux is added, 129 Kg of coal is used as fuel, and oxygen-enriched air with 60% oxygen is used to assist the gasification reaction. The air temperature is 25 °C, and the nitrogen content is 40%. 260 m of air is blown in 3 Measured by gas volume. The designed pre-reduction metallization rate of the raw materials is 70%.

[0204] (1) Heat income Q

[0205] A. Heat generated from the reaction of FeS with oxygen: -195539 KJ (3 - 2 - 1 - A)

[0206] B. Heat generated from the combustion of 129 kg of coal: -1177883.52 KJ (3 - 3 - 1 - B)

[0207] C. Heat generated from the reaction of N0 with Fe: -36139.6 KJ (3 - 2 - 3 - A)

[0208] D: Heat of reaction generated by CuO and Fe: -60622.14 KJ (3-2-4-A)

[0209] The total is: -1470184.26 KJ.

[0210] (3) Heat expenditure:

[0211] A: Endothermic heat during the reduction reaction of Fe3O4: 45266.59 KJ (3-2-2-A)

[0212] B: Endothermic heat during the reduction reaction of FeO: 34870.73 KJ (3-3-2-A-8)

[0213] C: Physical heat carried away after the reaction of the reducing gas: 355456.25 KJ (B-3-2-C-1)

[0214] D: Thermal energy required to melt 1000 kg of raw materials: 1076630 KJ (3-1-1)

[0215] (4) Heat balance (heat income + heat expenditure)

[0216] (-1470184.26) + 1512223.57 = 42039.31 KJ.

[0217] The heat difference of 42039.31 KJ / 3600 Kw.h = 11.7 Kw is supplemented by electric heating.

[0218] (5) Physical and chemical energy carried in the flue gas and tail gas

[0219] The physical energy in the tail gas is 355456.25 KJ.

[0220] The physical energy in the flue gas is 71091.25 KJ.

[0221] The total physical energy carried away is 426547.5 KJ.

[0222] The chemical energy is CO. The total is 2.187 + 2.175 = 4.362 Kmol.

[0223] 3.5 Calculation of the heat energy for rotary kiln calcination.

[0224] The temperature of the raw materials and combustion-supporting air is 25 °C. The raw materials are set to contain 35% water. Calculated based on 1000 kg of raw materials entering the electric furnace, the weight of the raw materials containing water is 1000 ÷ 65% = 1538.5 kg, and the water to be calcined and evaporated is 538.5 Kg.

[0225] A: 1000 Kg of raw materials come out of the kiln at a temperature of 350 °C with 0% water content, and the average heat enthalpy of the materials is 0.9362 KJ / Kg.K.

[0226] The calorific value carried away by the material is calculated as follows:

[0227] H = 1000 × 0.9362 × (330 °C - 25 °C) = 304265 KJ

[0228] B: The heat of vaporization of water and the weight of evaporated water is 538.5 Kg

[0229] H = 538.5 × 2255.2 = 1214425.2 KJ

[0230] C: The physical heat carried in by the flue gas is: 426547.5 KJ (3 - 4 - 5)

[0231] D: The chemical energy carried in by the tail gas and the flue gas.

[0232] According to the description in section (3 - 4 - 5), the amount of CO is 4.362 Kmol, and the exothermic reaction of CO burning with oxygen in the air is as follows:

[0233] CO + 0 = CO2 H f298 = -282946 KJ / KmoL

[0234] The heat released Q = 4.362 × (-282946) = -1234288.968 KJ.

[0235] The amount of oxygen required for the reaction is: 4.362 ÷ 2 = 2.187 Kmol.

[0236] Using air for combustion support, the air excess coefficient is 1.2

[0237] The volume of oxygen used is: 2.181 × 1.2 × 22.4 m 3 / Kmol = 58.6 m 3

[0238] The volume of air is 58.6 ÷ 21% = 279 m 3

[0239]

[0240] Table (8) Composition of the rotary kiln tail gas and the heat carried away (excluding the moisture dried in the raw materials)

[0241] Composition <![CDATA[CO2]]> <![CDATA[SO2]]> <![CDATA[H2O]]> <![CDATA[N2]]> Total Molar quantity (kmol) 10.062 0.418 2.457 13.20 26.137 Mole fraction % 38.5 1.6 9.4 50.5 100 Mass kg 442.7 26.76 44.23 369.6 883.3 <![CDATA[Standard state volume m 3 > 224 9.4 294 527.5 Average specific heat capacity 49.53 50.88 37.62 31.05 Unit heat capacity 498.37 21.27 92.43 409.86 1021.93

[0242] Due to the excessive moisture carried in the flue gas, the tail gas emission temperature is required to be controlled at about 150°C. The water dried out from the raw materials is not calculated in the above formula. The design relaxation of the water content is considered because the equipment may produce laterite ore (nickel) and hazardous waste disposal. When using other raw materials for production, such a high moisture content is impossible. If the raw material water truly has a water content of 35%, the mole fraction of H2O in the flue gas tail gas is 58%. This is very difficult for flue gas dust removal, etc. It is necessary to increase the flue gas flow rate to reduce the mole fraction of H2O in the flue gas. Therefore, in actual production, corresponding adjustments should be made according to the water content of the raw materials. So here the calculation does not include the water dried out from the raw materials. However, the flue gas design process leaves room for adjustment.

[0243] Due to the excessive moisture carried in the flue gas, the tail gas emission temperature is required to be controlled at about 150°C. The water dried out from the raw materials is not calculated in the above formula. The design relaxation of the water content is considered because the equipment may produce laterite ore (nickel) and hazardous waste disposal. When using other raw materials for production, such a high moisture content is impossible. If the raw material water truly has a water content of 35%, the mole fraction of H2O in the flue gas tail gas is 58%. This is very difficult for flue gas dust removal, etc. It is necessary to increase the flue gas flow rate to reduce the mole fraction of H2O in the flue gas. Therefore, in actual production, corresponding adjustments should be made according to the water content of the raw materials. So here the calculation does not include the water dried out from the raw materials. However, the flue gas design process leaves room for adjustment.

[0244] The heat carried away by the flue gas is: H = 1021.93×1.2×(150°C - 25°C) = 153289.5 KJ;

[0245] The standard state volume of the flue gas emission is: 527.5×1.2 = 633 m 3 ;

[0246] F: Heat balance analysis (rotary kiln part)

[0247] The heat income Q = (-426547.5) + (-1234288.968) = -1660837 Kg.

[0248] The heat expenditure H = 304265 + 1214425.2 + 153289.5 = 1671980 KJ.

[0249] The heat difference = -1660837 + 1671980 = 11143 KJ = 3.1 Kw.h.

[0250] IV: Equipment selection and design

[0251] 4.1: Production capacity

[0252] Taking the raw materials in Table (4) as an example, for pure electric production of 30,000 tons / year and auxiliary oxygen lance heating for pre-reduction production of 50,000 tons / year (note: measured by the weight of the raw materials), the equipment is selected and designed as follows:

[0253] The water content of the raw materials is 35%.

[0254] The monthly production capacity is: for pure electric production, 30,000 t / 300 days = 100 t / day;

[0255] For oxygen - coal assisted heating production = 50,000 t / 300 days = 167 t / day.

[0256] Since the peak electricity price is applicable from 19:00 to 21:00 in most areas. And the electric furnace also needs to be overhauled and maintained. At this time, it just takes advantage of the peak - electricity period for equipment overhaul and maintenance. So we calculate based on 22 - hour daily production:

[0257] The hourly production capacity is:

[0258] For pure electric production, it is 100 t / 22 h = 4.55 t / h;

[0259] For oxygen - coal assisted heating production, it is 167 t / 22 h = 7.6 t / h.

[0260] Since the raw materials contain 35% water. About 20% of coal and flux need to be added to the raw materials.

[0261] The input amount into the furnace for pure electric production is: 4.55×(1 - 35%)×12% = 3.55 t / h;

[0262] The production amount for oxygen - coal assisted heating is: 7.6×(1 - 35%)×12% = 5.93 t / h.

[0263] 4.2 Selection of the calcining rotary kiln

[0264] According to the calculation in subsection (4.1), the production amount of oxygen - coal assisted heating. The raw materials are 7.6 t / h, with a water content of 35%. The temperature of the raw materials entering the kiln is 25 °C, the temperature leaving the kiln is 350 °C, the temperature of the flue gas entering the kiln is 1100 °C, and the temperature of the flue gas discharged from the kiln is about 150 °C. The material filling rate is about 20%, and the filling angle is 120°. The residence time of the material in the kiln is not less than 3 hours. Since the installation slope of the rotary kiln is 2.5 - 3°, and the specific gravity of the raw materials is 1.5 t / m 3 The maximum rotational speed of the rotary kiln is not more than 3 revolutions per minute.

[0265] Based on the conditions provided above, we select the rotational speed of the rotary kiln during normal production to be 2.6 revolutions per minute and the filling angle to be 120°. It means that the material moves forward 3 times per revolution of the rotary kiln. We set the forward - movement distance each time to be 25 mm.

[0266] The required length of the rotary kiln is: 25 mm×3 times×2.6 revolutions per minute×180 minutes = 35.1 m.

[0267] So we select the length of the rotary kiln to be 36 m.

[0268] Since about 20% of coal and flux need to be added to the materials before they enter the rotary kiln.

[0269] The actual amount of wet-based materials entering the kiln is: 7.6 + (7.6 × 65% × 20%) ≈ 8.6 t / h.

[0270] The volume of materials entering the kiln per minute is: 8600 kg / h ÷ 1500 Kg / m 3 ÷ 60 min = 0.096 m 3 / min.

[0271] The moving distance of materials in the kiln per minute is: 25 mm × 3 times × 26 revolutions / min = 0.195 m / min.

[0272] Since the filling rate of materials in the rotary kiln is 20%.

[0273] The inner diameter of the rotary kiln is 0.096 m 3 / min ÷ 0.195 m 2 / min ÷ 20% = 2.46 m 2 .

[0274] Therefore, the inner diameter of the rotary kiln is: 1.77 m.

[0275] We choose the inner diameter of the rotary kiln to be 1.8 m and the single-sided refractory material to be 150 mm. That is, the outer diameter of the rotary kiln is 2100 mm. The materials are pushed forward 25 mm each time. Therefore, the installation slope of the rotary kiln body is selected to be 2.5°;

[0276] The working cross-sectional area of the rotary kiln is: (0.9) 2 × 3.14 = 2.54 m 2 .

[0277] The production volume under ideal conditions is: 2.54 m / s × 20% × 0.195 m / min × 1500 Kg / m 3 × 60 min = 8915 kg / h.

[0278] 8915 kg / h is greater than the actual amount of wet-based materials entering the kiln. It shows that the rotary kiln can meet the normal production demand.

[0279] According to the calculation in Table (7), the flue gas entering the rotary kiln is 633 m 3 / t. The current designed production volume is 5 t / h. Based on the amount of materials entering the electric furnace, that is, the amount of flue gas entering the rotary kiln is: 633 m 3 / t × 5 t / h = 3165 m 3 / h.

[0280] The flue gas passes through the rotary kiln, and the wind speed in the kiln is 3165 m 3 / h÷(2.54×80%)÷3600 = 0.44 m / s.

[0281] 4.3 Selection of Melting Electric Furnace

[0282] 4.3.1 Selection of Transformer Power of Melting Electric Furnace

[0283] A. According to the existing data, find out the power consumption per ton of raw materials, and then select the transformer capacity. The current pure electric production is 3 tons per hour. Based on the existing table (1) materials. According to the data, calculated by the raw material charge into the furnace: the power consumption of hot charging melting of materials is 600 kw.h / t. So the transformer power should be 1800 kw.

[0284] B. According to the calculation in Section 3.4.3, the heat energy required for material reaction and melting is 1512223.57 KJ,

[0285] That is, 420 kw.h / t. The heat loss and other losses of the electric furnace are all 130 kw.h / t. The total required energy power is 550 kw.h / t.

[0286] Based on the above data, for pure electric production with a production volume of 3 t / h. The power of the electric furnace transformer should be selected as 1800 kw, the transformer impedance is 8%, so the transformer capacity is 1950 RVA. Adopt strong oil-water cooling, and the transformer has the ability to be overloaded by 20% for a short time. It can withstand short-circuit impact.

[0287] 4.3.2 Working Voltage

[0288] This project adopts the design of a DC electric furnace. Due to the lack of design data for DC electric furnaces. We use the formula of the design data of three-phase alternating current to calculate the DC working voltage. Finally, verify it by other methods.

[0289] According to the empirical formula of three-phase AC electric furnace design. The secondary rated voltage of the furnace transformer is: U 线 = KP n where K is the empirical coefficient: K = 13.5, n is the production coefficient of nickel oxide ore: n = 0.33. So U 线 = 13.5×1800 0.33 km / m33 = 164.16 V. The phase voltage of the electrode to the furnace bottom is 1164.16÷1.73 = 94.89 V. So we select the rated working DC voltage of the electric furnace power supply to be 100 V. To adapt to the production of various raw materials, we require the DC power supply to be adjustable to: 1. 52 V. 2. 60 V. 3. 68 V. 4. 76 V. 5. 84 V. 6. 92 V. 7. 100 V. 8. 108 V. 9. 116 V. 10. 124 V. 11. 132 V. Among them, 1-6 are constant current, and 7-11 are constant power.

[0290] 4.3.3 Rated Working Current

[0291] Rated working current: Ⅰ = P / U = 1800 / 100 = 18000 A

[0292] 4.3.4 Electrode Design

[0293] A: The minimum cross-sectional area of the electrode is 18000 A / 6 A / cm 2 = 3000 cm 2 . (Note: The recommended allowable current density for the consumable electrode is 6 A / m 2 ).

[0294] B: According to Table (7), when the auxiliary oxygen lance is used for heating and blowing, the standard volume of the reducing gas generated is 260 m 3 / t.

[0295] The designed production capacity is 5 t / h, so the amount of the standard-state reducing gas generated during production is: 260 × 5 = 1300 m 3 / h.

[0296] That is, 1300 ÷ 3600 = 0.36 m 3 / s.

[0297] Requirements: The porosity of the material in the hollow electrode is about 0.3, and the wind speed is not greater than 3 m / s. Therefore, the inner diameter of the hollow electrode is: 0.36 ÷ 0.3 ÷ 3 = 0.4 m 2 , that is, the inner diameter of the hollow electrode is 0.72 m.

[0298] Based on the above calculated data and requirements, we select a hollow electrode with an inner diameter of 0.75 m (750 cm) and an electrode wall thickness of 12.5 cm. That is, the outer diameter of the electrode is 100 cm (1 m). The cross-sectional area of the electrode is: 3434 cm 2 ; The cross-sectional area of the electrode is 3434 cm 2 ; which is larger than the minimum cross-sectional area of 3000 cm required for the electrode to carry current 2 . In fact, the recommended allowable current density of the DC electrode can reach 10 A / cm 2 or more. As long as the strength of the hollow electrode is sufficient, there is no problem with the electrode.

[0299] C: Verification voltage: The conductivity of the general electric furnace slag is 2 Ω.cm, that is, f = 2 Ω.cm. Calculation method of the circumferential resistance coefficient.

[0300] The resistance of the electrode end to the furnace bottom is: R = f / πd = 2 / 3.14 * 100 = 0.00637 Ω;

[0301] The voltage of the electrode end to the furnace bottom is: U phase = 107 V.

[0302] According to the calculation in Subsection 4.3.2, the rated voltage is 100V, and the power supply voltage we selected is adjustable. The design data can meet the requirements of actual production.

[0303] D. Electrode consumption

[0304] According to A: The minimum cross-sectional area of the electrode is 3000 cm 2 , that is, the diameter of the solid electrode is d = 62 cm; the end face of the electrode is spherical, and the hemispherical area is 6035 cm 2 ; Suppose the insertion depth into the slag is 40 cm, and the spherical radius of the electrode end face is 31 cm. So the cylindrical surface of the electrode immersed in the slag is: 40 cm - 31 cm.

[0305] The area of the electrode cylindrical surface immersed in the slag is: 1947 cm 2 ;

[0306] The total area of the solid electrode immersed in the slag is: 7982 cm 2 ;

[0307] Calculation of the area of the hollow electrode immersed in the slag. The depth of the electrode immersed in the slag is 40 cm;

[0308] The end face of the electrode is spherical-arc-shaped, and the arc area is: 5393 cm 2 ;

[0309] The area of the electrode cylindrical surface immersed in the slag is: 18546 cm 2 ;

[0310] The total area of the hollow electrode immersed in the slag is: 23939 cm 2 .

[0311] The area ratio of the hollow electrode immersed in the slag to the hollow electrode immersed area is 23939÷7982≈3 times.

[0312] The above data shows that because the inner and outer surfaces of the hollow electrode are in contact with the slag surface, the contact area is relatively large, the current passing through is relatively large, and the consumption is also correspondingly large. Therefore, in actual work, the working voltage should be smaller than the theoretical calculated value. In actual production, the electrode consumption is about 2 times that of the solid electrode. Attention should be paid when designing the electrode.

[0313] According to experience, when producing nickel-copper alloy, the consumption of electrode paste is 8 g / Kw.h. When we produce at the maximum power, the consumption of electrode paste is: 691.2 kg / day, and the apparent density of the electrode paste is 1450 kg / m 2 , so 6912÷1450 kg / m 3 = 0.477 m 3 .

[0314] With the total area of the hollow electrode end face being 3434 cm 2, the daily consumption length of the hollow electrode is calculated as follows:

[0315] 0.477m 2 ÷(3434cm 2 ÷10000) = 1.389m.

[0316] Therefore, it is required that the monthly baking length of the hollow self-baking electrode is 1.6m. Since the daily baking length of other solid electrodes does not exceed 0.5 meters. Therefore, special attention should be paid when ordering the equipment, and it must be ensured that the hollow electrode from the head can be used normally.

[0317] 4.3.5 Furnace diameter

[0318] The furnace diameter is D = d + 2×cd

[0319] where d is the diameter of the hollow electrode, c is the empirical coefficient c = 1.2, so D = 3400mm.

[0320] 4.3.6 Total furnace height

[0321] The total furnace height H = h_gold + h_slag + h_material + h_gas

[0322] In the formula, the height of the metal: h_gold = 300mm, the height of the slag layer: h_slag = 500 - 600mm, the height of the melted material

[0323] layer height: h_material = 400 - 500mm, the height of the gas space: h_gas = 600mm.

[0324] Since there is space in the furnace cover, with a height of about 400 - 500mm, the gas height in the furnace here is taken as 200mm.

[0325] So H = 300 + 500 + 500 + 200 = 1500mm.

[0326] 4.3.7 Power density

[0327] A: Area power density of the reaction zone

[0328] Since the hollow DC electrode melting is adopted, and the melting zone is inside the electrode. When melting materials with heavy electricity, the area inside the electrode is S = 0.785m 2 , and the area power density is: 2293kw / m 2 .

[0329] B: Volume power density of the reaction zone

[0330] The power density of the method adopted in this project is very high, and the calculation is meaningless. However, considering that the electrode may feed and melt raw materials for smelting in a general way, that is, melting and smelting with external feeding of the electrode. Therefore, the conventional method is used for calculation and the original volume power density should be used. The volume power of the reaction zone is 396Kw / m 3。The electric furnace can also be applicable to the production of nickel chromium iron by other smelting methods.

[0331] C: Hearth area power density

[0332] The hearth area is S = 9m 2 , and the power density is 200 kw / m 2 . The amount of metal layer is: 13.5t, and the amount of slag is: 15.3t.

[0333] D: Hearth volume rate density

[0334] The hearth volume is: 13.5m 3 , and the volume power density is: 133.3Kw / m 3 .

[0335] 4.3.8 Low smoke hood

[0336] Adopting a low smoke hood design can reduce the amount of flue gas and heat loss. The smoke hood can adopt a water-cooled beam support and a high-temperature resistant concrete casting structure. It is necessary to reserve a flue gas discharge port, a feeding hole, an operation port, and a safety explosion-proof hole.

[0337] 4.3.9 Combined auxiliary oxygen lance

[0338] According to the requirements of this project, air is transported through the outer pipeline of the central oxygen delivery pipeline. Normal temperature air is used to cool the oxygen lance and protect the oxygen lance. At the same time, the outer transported air can also be used for transporting pulverized coal or other fuels. Calculated based on the data in Table (6): The designed transportation flow rate of the central oxygen pipe is 120m 3 / h; the mixing of the outer air transportation pipeline is 150m 3 / h; the air flow velocity should not be greater than 60m / s; the oxygen transportation should not be greater than 180m / s. Ensure the pressure drop and safety of gas transportation.

[0339] The local temperature at which the oxygen lance works is 1000°C. Refractory bricks can be installed outside the oxygen lance for protection. Ensure safe use and extend the service life. Two oxygen lances are adopted, one is in use and the other is in reserve to ensure normal use and safety. The oxygen lance should be regularly overhauled and maintained. It must be replaced in time when the service life is reached.

[0340] 4.3.10 Foundation, support, platform, flue

[0341] When designing the foundation, the heaviness of the furnace equipment should be fully considered. At the same time, it should also be considered that in case of an accident, the metal and slag leaking from the furnace can be drained into the accident pit. The size and capacity of the accident pit should be appropriate to accommodate the metal and slag leaking during an accident, and should not affect the safety of the furnace foundation and platform. Minimize the damage to the foundation and equipment caused by the accident.

[0342] The slag tapping hole and the iron tapping hole of the electric furnace should be reasonably arranged for convenient and safe operation. There should be sufficient space for heat dissipation. Generally, the slag tapping hole and the iron tapping hole are in two different positions without interference. When the iron tapping hole and the slag tapping hole are the same, a spare outlet should be reserved on the other side of the electric furnace to ensure its use in case of emergency. A smoke hood should be made above the iron tapping hole and the slag tapping hole. When tapping iron and slag, the flue gas can be sucked away to ensure environmentally friendly and safe operation.

[0343] The foundation should reserve the cable pipeline and magnetic shielding for the electric furnace. Since the magnetic field generated by the DC circuit cable will cause arc magnetic blowout, affecting the melting of materials and the impact of the arc on the furnace wall, which is also one of the defects of the DC arc furnace, attention should be paid during the design.

[0344] The platform design from bottom to top is respectively: (1) the iron and slag tapping layer; (2) the operation platform layer; (3) the feeding platform layer. The platform is welded by iron brackets and iron plates. The operation platform of the operation platform layer can be poured with concrete on the iron plate for heat insulation to improve the operation environment.

[0345] Since the electrode moves up and down, a flue gas seal for the up and down movement of the electrode needs to be designed. The interface should ensure that a large amount of high-temperature flue gas generated by the oxygen lance auxiliary heating can be discharged in time with low resistance without affecting the up and down movement of the electrode. At the same time, considering the large dust content in the flue gas, there should be a suitable space for dust settlement and recycling. The space between the upper part of the (2)-layer electric furnace and the (3)-layer platform can be separated. Without affecting the normal operation of smelting, it can also be used for the production of self-baking electrodes to ensure the safe operation of the self-baking electrode production. The space of the (3)-layer platform is mainly for the storage bin of high-temperature materials and the distribution and quantitative transportation into the electric furnace.

[0346] Based on the above data conditions, a suitable manufacturer can be selected for ordering and equipment design.

[0347] Any embodiment of the present application can be used as an independent technical solution or can be combined with other embodiments. All patents and publications mentioned in the specification of the present application indicate that these are publicly known technologies in the art and can be used in the present application. All patents and publications cited herein are equally listed in the references, just as each publication is specifically individually referenced. The present application herein can be implemented in the absence of any one or more elements, one or more limitations, where such limitations are not specifically stated. The terms and expressions used herein are for the purpose of description and are not limiting, and there is no intention herein to indicate that these terms and interpretations described herein exclude any equivalent features, but it can be understood that any suitable changes or modifications can be made within the scope of the present application and the claims. It can be understood that the embodiments described in the present application are some examples and features in some embodiments, and any person of ordinary skill in the art can make some changes and variations based on the essence described in the present application, and these changes and variations are also considered to be within the scope of the present application and the scope limited by the independent claims and the dependent claims.

Claims

1. A metallurgical production process using a hollow electrode, characterized in that, It includes the following steps: 1) Calcining and drying the raw materials; 2) Pre-reducing and melting; Feeding the calcined and dried raw materials into the hollow electrode. The raw materials freely fall from the upper part of the hollow electrode under the action of their own weight. The oxygen-rich gas and fuel are blown out by the combined oxygen lance located at the center of the electrode. The oxygen-rich gas reacts with carbon, hydrogen and metal oxides in the raw materials and fuel to release heat and generate reducing gas. The released heat melts the raw materials, and the reducing gas contacts the raw materials countercurrently from bottom to top to reduce the oxides in the raw materials and generate metals; 3) Continuously adding materials. After the materials are melted, coal is concentrated, and an annular solid phase zone is formed in the hollow electrode part. The reduced metals fall under the action of their own weight and melt into a liquid state. After the liquid metal or metal slag passes through the solid phase zone, it precipitates into the bottom metal layer. When the remaining metal oxides pass through the solid layer, they continue to react with carbon in the solid phase zone to generate metals; Among them, in step 2, the hollow electrode is inserted into the slag surface, and a positive pressure is formed inside the hollow electrode by using the slag liquid seal. The combined oxygen lance includes a central oxygen delivery pipe, an air pipe, a fuel pipe and a forced water cooling pipe, or the combined oxygen lance is composed of a central oxygen delivery pipe and an outer air cooling pipe body.

2. The metallurgical production process of the hollow electrode according to claim 1, characterized in that: In step 1, the raw materials and coal particles are calcined and dried. In step 2, the calcined and dried raw materials and coal particles are fed into the hollow electrode. The oxygen-rich gas is blown out by the combined oxygen lance located at the center of the electrode. The oxygen-rich gas reacts with carbon, hydrogen and metal oxides in the raw materials and coal particles to release heat and generate reducing gas.

3. The metallurgical production process of the hollow electrode according to claim 2, characterized in that: In step 1, the raw materials, fluxes and coal particles are calcined and dried. In step 2, the calcined and dried raw materials, fluxes and coal particles are fed into the hollow electrode.

4. The metallurgical production process of the hollow electrode according to any one of claims 1-3, characterized in that: Step 1 is specifically as follows: The raw materials are fed into the rotary kiln by a conveyor. The rotary kiln calcines the raw materials in a countercurrent manner. The required temperature of the discharged flue gas is ≤168°C. The water content of the calcined materials is ≤2%, and the remaining temperature of the calcined materials is 350°C ± 10°C.

5. The metallurgical production process of the hollow electrode according to claim 4, characterized in that: The raw materials are transported and lifted by a belt conveyor and fed into the rotary kiln by a screw conveyor. The rotary kiln calcines the raw materials in a countercurrent manner. The rotation speed of the rotary kiln is 1 - 3 revolutions per minute. The materials stay in the rotary kiln for no less than 3 hours. The head of the kiln is high for feeding materials, and the tail of the kiln is low for discharging materials. The required temperature of the flue gas discharged from the head of the kiln is ≤168°C, which is convenient for using reliable and economically priced filter bags; the water content of the calcined materials is ≤2%, and the remaining temperature of the calcined materials is 350°C ± 10°C.

6. The metallurgical production process of the hollow electrode according to claim 1, characterized in that: The heat for calcining in step 1 comes from the physical heat in the pre-reduced flue gas in step 2 and the heat generated by secondary combustion. CO and H2 contained in the flue gas are burned in the secondary fuel chamber added at the tail of the kiln before the flue gas enters the rotary kiln.

7. The metallurgical production process of the hollow electrode according to claim 1, characterized in that: The combined oxygen lance is installed and aligned with the center point of the hollow electrode. The combined oxygen lance is insulated from the electrode during installation and use. The tip of the combined oxygen lance is 500 - 800 mm away from the slag surface.

8. A metallurgical electric furnace, characterized in that: The metallurgical electric furnace is applied to the hollow electrode metallurgical production process described in any one of claims 1 to 7, and includes a furnace wall (1) and a furnace cover (2). A hollow electrode (3) is installed on the furnace cover (2), and a combined oxygen lance (4) is installed at the center point of the hollow electrode (3). The combined oxygen lance (4) includes a material pipe and a cooling pipe. The material pipe includes a central oxygen delivery pipe, an air pipe, and a fuel pipe. The cooling pipe includes a forced water cooling pipe or an outer layer air cooling pipe.

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