RU2228362C2 - Method of blast-furnace smelting - Google Patents

Abstract

FIELD: metallurgy; method of blast-furnace smelting. SUBSTANCE: proposed method includes separate loading of coke and iron ores through furnace top, introduction of additional carbon-containing additive into iron ore material in the amount of 10- 100 kg/ton-h above total consumption of fuel through furnace top and tuyeres. Used as additives are coke wastes, crushed anthracite, granulated coal at size not exceeding 5 mm and water-and- coat slime. Degree of homogeneity of iron ore materials and carbon-containing additive does not exceed 2. Average size of iron ore materials does not exceed 40 mm and size of blast-furnace coke is no less than 40 mm. Smooth travel coefficient which is no less than 0.5 is used for check of stability of smelting process. EFFECT: enhanced gas permeability of ore softening zone. 5 cl, 18 dwg,13 tbl,2 ex

Description

The invention relates to ferrous metallurgy, in particular to blast furnace production.

A known method of blast furnace smelting, including loading all coke and all ore onto the top in a mixed state with a grain size of materials within 10-15 mm. The disadvantage of this method is the reduction of gas permeability of the column of materials, the violation of the stability of the movement of the column [1].

The closest analogue is the method of blast furnace smelting, which includes separate loading through the top of iron ore materials and coke by feeds, loading of a carbon-containing additive, blowing through lances of blast and coal dust [2].

The disadvantage of this method is the impossibility of smelting small ore materials, the use of increased costs of coal dust in the blast.

The object of the invention is to ensure that the stroke is within acceptable limits when smelting small iron ore materials and increased consumption of coal dust in the blast.

The technical result that is achieved in the invention is to increase the gas permeability of the softening zone of iron ore materials in a mixture with a carbon-containing additive.

To achieve a technical result in a blast furnace smelting process, comprising separately loading iron ore materials and coke through a top, feeding a carbon-containing additive, blowing blast and coal dust into the tuyeres, according to the invention, iron ore materials are mixed with a carbon-containing additive in an amount of 10-100 kg before being loaded onto the furnace top. / t of cast iron in excess of the total fuel consumption, loaded through the top and blown into the tuyeres, while the degree of homogeneity of the mixture of iron ore materials and carbon-containing the allowance is no more than two, and the particle size of the carbon-containing additive does not exceed 5 mm.

Moreover, as a carbon-containing additive, coke waste, crushed anthracite, granule, and coal-water suspension are used.

The average size of iron ore materials is not more than 40 mm, the size of coke is not less than 40 mm.

In this case, the evenness coefficient of the furnace is determined and when the porosity of iron ore materials is below 0.3 m ³ / m ^{3, the} consumption of carbon-containing additives is changed in doses of 10 kg / t of pig iron for every 0.01 deviation of the evenness coefficient below the permissible level of 0.5.

At the same time, with a consumption of coal dust blown into the tuyeres within 0.2-0.4 t / t of pig iron, the consumption of carbon-containing additive is 50-100 kg / t of cast iron, while the supply of coke and iron ore materials is loaded through the top with compact layers within 1-5 innings in each layer.

The invention is illustrated in figures 1-18, which depict the following.

Figure 1. Scheme of the loading path according to the method.

Figure 2. Scheme of horizontal softening volume one meter high, equal to the real volume (arbitrary scale).

The location of the horizontal volume is specified by the condition that their centers of gravity coincide.

Figure 3 a. Shrinkage curves of ore material in the softening volume, carbon loss of the additive according to the Boudoir reaction, position of the calculated volume over the furnace height depending on the gas temperature and the initial softening temperature of the ore substance.

Figure 3 b. The volumetric scheme of shrinkage of the ore column with a base area of 1 cm ² (arbitrary scale).

Figure 4 a. Dependence of the evenness coefficient and the degree of variability of the hot blast pressure curve with a decrease in the gas permeability of the softening plate material during smelting of fine ore due to an increase in the consumption of coal dust over 0.2 t / t of cast iron.

Figure 4 b. Zet-nomogram of the relationship between the coefficients of evenness and unevenness (unevenness) of the course.

Figure 5. The dependence of the specific surface of the granules of the additive on the diameter at different consumption of the additive. The ratio of the end time of the Boudoir reaction in the plate to the residence time of the plate in the fixed volume of the furnace in which softening processes take place.

6. Scheme of proper packing of four ore balls of equal diameter in the dry zone of the furnace.

7. Dependence of the diameter of the through channels in the dry ore layer on the diameter of the balls.

Fig. 8. Dependence of the diameter of the shell of the granule of the additive in the plate reduced to a continuous ball on the diameter of the granule.

Fig.9. Diagram of a diametrical section of a granule shell with a granule diameter of 3 mm.

Figure 10. The load balance on the diametrical section of the shell of the granule of the additive in the plate.

11. The dependence of the geometric increase in the porosity of the plate after burning the granules of the additive on the diameter of the granule and the degree of use of the substance of the additive according to the Boudoir reaction with the consumption of the additive in the range of 50-100 kg / th (table 7).

Fig. 12. The balance of loads on the plate, depending on the test porosity of the substance of the plate according to the first embodiment of the method (smelting of fine ore).

Fig.13. Dependence of ore load on residual coke on the consumption of coal dust introduced with blast through tuyeres.

Fig.14. The effect of ore shrinkage on the porosity of the ore substance in the softening volume.

Fig.15. The dependence of the coefficient of gravitational load on the plate from the test porosity of the plate material during the melting of fine ore, the consumption of coal dust in the range 0-0.4 t / t h

Fig.16. The balance of the available increase in porosity after burning away the granules of the additive and the necessary increase in porosity to create a positive balance of loads on the plate by the method.

Fig.17. Diagram of thermal stability of the heat in the graphical interpretation of ROSA.

Fig. 18. Porosity curves of ore and coke depending on the distance above the tuyere level.

Figure 1 shows a diagram of the loading path to the top according to the method in which: 1 - plan of the top, 2 - receiving hoppers of the filling apparatus, 3 - coke supply line, 4 - screening of fine fractions below 40 mm, 5 - sorted coke, 6 - crusher, 7 - screening of small fractions of coke above 5 mm with a return line to the crusher, 8 - finished additive; 9 - feed line of components of ore materials with a grain size of not more than 40 mm, 10 - mixer, 11 - feed line of solid carbon-containing additives, 12 - feed component of the ore mixture with the additive at the top, 13 - feed line of coal-water slurry (sludge).

The scheme works as follows. The mixer 10 receives components of iron ore materials (agglomerate, etc.) with a grain size of not more than (average) 40 mm, solid carbon-containing additives with a particle size of not more than 5 mm, a water-carbon suspension with inclusions of solid particles with a particle size of not more than 5 mm. After mixing, the finished mixture — an ore product with an additive — enters the receiving hopper 2 on the furnace top. Sorted coke with a fineness (on average) of at least 40 mm comes there. Both components from the receiving hoppers 2 separately enter the top and are placed over the cross-sectional area of the top in the accepted stacking order. If there is a shortage of fine coke, coarse coke screenings are added to the crusher 6. The mode of operation of the mixer 9 provides the production of a finished ore product charged with an additive with a mixture homogeneity coefficient of not more than two (the ratio of the concentration of the additive in the samples to the average concentration of the additive over the entire mass of the finished ore product) [3].

There are two options for using the method: 1) to compensate for violations of the movement of the column during the smelting of small ores; 2) to compensate for furnace malfunctions with increased consumption of coal dust in the blast to the level of 0.4 t / t h. The consumption of dry solid additive (or reduced to the dry residue of a coal-water suspension) is considered to be in excess of the accepted consumption of the amount of coke (K) plus coal dust ( S _yy ). Conditional fuel consumption T = K + S _ug t / t h. The coefficient of coke replacement for coal dust was adopted for the content of non-volatile carbon equal to E = 1. For other replacement factors, the load balances in the softening zone should be recalculated using the formulas below. It is accepted that the column motion is determined by the gas permeability of the softening volume. In this case, the column motion and the motion of the softening layer will be identical according to the equation of continuity of flow [5]. The additive carbon partially replaces the coke carbon consumed by the Boudoir reaction in the softening volume, providing a decrease in the coke consumption below the calculation. The introduction of additives does not affect the porosity of coke over the entire height of the column of materials, which clearly depends on the grinding of top coke when lowering the composition of the column of materials.

For the second variant of the method, the complete combustion of the increased consumption of coal dust inside the volume of the tuyere cavity due to enrichment of the blast with oxygen and increase in the volume of the tuyere cavity with a coefficient of evenness close to 0.5 is accepted. The complete combustion of coal dust in the volume of the tuyere cavity eliminates the saturation of the liquid smelting products with unburned dust residues and facilitates the release of melts from the notches. The thermal state of the furnace is considered unchanged, provided by the control of the observed parameters of gas use and the composition of cast iron released from the notch.

B. Thermomechanics of the process of softening ore.

When the slag exit is not more than 0.4 t / t h, the porosity of the layer in the melting and drainage zone, including liquid melt and a coke nozzle, is in the range of 0.2-0.4 m ³ / m ³ [4]. In the overlying softening volume, the porosity of the ore layer is approximately half that of the porosity of the drainage layer. Due to the shrinkage of the ore in the softening volume, hardened plates (cork) are formed, overlapping the furnace section. The strength of such plates increases with an increase in the fraction of the fine ore fraction or an increase in the volume fraction of ore inside the furnace when coal dust is introduced into the blast, as well as from fluctuations in the thermal state of the furnace [5]. For these reasons, the softening zone determines the movement of the column of materials.

Figure 2 shows a diagram of the horizontal softening volume equal to the real volume, and the positions of their centers of gravity along the height of the furnace coincide. The process inside the softening volume does not depend on the shape of the volume with the equality of the real and calculated volumes and the positions of their centers of gravity along the height of the furnace. Conventionally, we consider the height of the calculated layer in the softening zone to be equal to one meter, the diameter of the zone equal to the hydraulic diameter of the furnace. Ore and coke from separate layers laid on the top in the softening volume are collected in two compact volumes, which does not affect the calculation results provided that the sum of the volumes of the real layers of coke and ore and the calculated volume of one meter high are equal.

In figure 2, position 1 is the wall of the furnace; 2 - a central hole filled with large coke; 3 - the upper boundary of the volume; 4 - lower boundary; 5 - ore compact plate; 6 - the interface between the compacts of ore and coke within the estimated volume. Conventionally, we believe that the plate 5 covers the entire cross-sectional area of the furnace evenly. The effect of the hole 2 is taken into account by a decrease in the height of the plate. We believe that the distance of the softening volume from the tuyere level is unchanged. The deviation of the volume form from the real one does not affect the load balance in the softening zone. We believe that the introduction of the additive granules into the plate does not affect the volume and quantity of the substance of the plate due to the burning of granules when the plate moves down. According to [6], inside the plate 5, the volume of unmelted ore is taken equal to 0.5. The volume of slag-forming - 0.5 of the total volume of the plate. When the plate moves downward, the slag-forming strength decreases due to an increase in the temperature of the plate substance and a decrease in their viscosity.

We consider according to the method when introducing carbon-containing additives into ore materials, the Boudoir reaction time does not exceed the residence time of the plate in the softening volume according to the equation:

With _add + CO ₂ = 2CO.

Figure 3a shows curves 1 and 2 — the residual ore volume after shrinkage, depending on the temperature of the substance and the softening temperature in the softening volume (Fig. 2) (both temperature scales are aligned along the abscissa axis). Curve 3 - the decrease in the substance of the granules according to the Boudoir reaction. All values are given in relative units (along the ordinate axis). When the furnace is operated on one coke, the Boudoir reaction occurs partially in the dry zone, and partially in the melting and drainage zones. According to the method, the place of the Boudoir reaction from the melting and drainage zones is partially transferred to the volume of the ore plate. The position of curve 3 inside the plate also depends on the degree of graphitization of the carbon additive, porosity and reactivity of the granules. Due to the partial transfer of the Boudoir reaction from the drainage zone to the plate volume, foaming of liquid melts in the drainage zone and the likelihood of slag flooding are reduced. For the calculations, the position of the volume of the furnace in which the softening processes take place is assumed to be stationary, the initial softening temperature, and the position of the gas isotherms along the height of the furnace are unchanged.

Figure 3b shows the geometry of the ore column after shrinkage with a base of 1 cm ² . The height of the column is equal to the height of the ore layer on the top, the volume of which will be

V _count = V _mind + V _pl .

From where the shrinkage coefficient in the plate is equal to K _mind = V _mind / V _count .

On figa schematically shows the effect of porosity of the plate in one cubic meter of the volume of the plate according to the expression (1-ε _PL ) on the indicators of movement of the column according to the method. Curve 1 - the coefficient of evenness. Details of the calculation of this coefficient are given in [5]. Curve 2 - the lower acceptable limit of the coefficient of evenness, close to 0.5. Curve 3 - the effect of input additives on the coefficient of flatness. Curve 4 is the starting point of impaired column movement. Curve 5 - the oscillation of the pressure curve of the hot blast in relation to its projection to the length of the tortuous curve.

On figa - zetomogram connection coefficients of the paired parameters of the movement of the column. Line 1 - percentage scale of the coefficient of evenness (ρ _x ). Line 2 - the coefficient of unevenness of the course (δ _srab ). Line 3 is the sum of both coefficients equal to 100%. For any values of one coefficient, the second is 100% minus the first coefficient. Line 4 - the boundary of the stability of the mechanical operation of the furnace. Region A - stability at ρ _x > 50%. Region B is instability and mode impacts are required to bring the furnace to region A (changing the radius of the hole 2 in FIG. 2, creating a loose ring of type 9 on the furnace top, introducing additives by the method, etc.). The calculation of the indicators of the movement of the column is carried out using a computer program.

B. Parameters of the Boudoir reaction for an isolated granule surrounded by a plastic shell.

The introduction of a loosening additive increases the porosity of the plate, due to this it ensures destruction of the body of the plate, creates a positive balance of loads on the plate, which leads to a stable downward movement identical to the movement of the entire column of materials. The porosity of the coke layer in the softening volume does not affect the addition of the additive. To calculate the parameters of the Boudoir reaction, we accept the following initial data: a basic furnace with a volume of 2000 m ³ , a top diameter of 7.3 m, a working height (from the tuyeres to the level of the mound) - 20 m, productivity - 4000 t / day, blast consumption - 3500 m ³ / min, the total pressure drop of the gas is 1.5 kg / cm ² , the oxygen content in the blast is in the range of 21-60 (percent). Ore loading equal to 3. The equivalent fuel consumption is 0.5 t / t h, including the consumption of coal dust from zero to 0.4 t / t h. Additive consumption (in excess of the standard fuel) is in the range of 10-100 kg / t h

We believe that the consumption of the additive completely covers the carbon demand for the Boudoir reaction inside the plate. Charge parameters: ore consumption - 1.5 t / t h, coke consumption (base) - 0.5 t / t h. Bulk weights: ore - 1.5 t / m ³ , coke - 0.5 t / m ³ . The ore shrinkage coefficient is 0.4 of the ore volume at the top. The volume of residual ore in the plate V _PL = 1.5 / 1.5 (1-0.4) = 0.6 m ³ / t h. This volume does not depend on the consumption of coal dust. Porosity of coke is taken equal to 0.5 m ³ / m ³ (grinding of coke is not taken into account in the calculation). The porosity of the ore at the top was taken equal to 0.3 m ³ / m ³ (fineness not more than 40 mm).

When the granules of the additive burn out inside the plate, an increase in porosity is created, which increases the full porosity of the plate after shrinkage. Calculation of the effect of the additive on the load balance of the plate includes: 1) determination of the optimal granule size; 2) determination of the gravitational forces applied to the plate from above, taking into account the influence of the furnace walls; 3) determination of the pressure gradient of the gas applied to the plate from below; 4) determination of the neutral point at which the calculated loads on the plate correspond to the stable movement of the column in a production environment; 5) the effect of the additive on the destruction of the body of the plate and a decrease in the gas pressure gradient along the height of the plate with two variants of the method: smelting of fine ore and increased consumption of coal dust in the blast.

The granule size is determined by the conditions of destruction of the slag-forming medium of the plate due to the creation of gas inside the pore pressure during the burning out of the substance of the granules, retention of the granules of the additive in the mixture with the ore layer on the top, and ensuring maximum reactivity of the granules in terms of the specific surface area of the granules reduced to the volume of the ball.

Table 1 shows the data of the specific surface area of an isolated granule in the form of a ball on the diameter of the granule: S _d = 6 / d cm ² / cm ³ , as well as the specific surface of the sum of balls with a diameter equivalent in volume to one ball with a diameter of 0.5 cm according to the balance of volumes (the specific gravity of the balls is the same, equal to 0.001 kg / cm ³ ). N _equiv V _d = V _0.5 . Whence the equivalent number of balls of a given diameter N _d = (0.5 / d) ³ . The total area of equivalent balls S _equiv = N _d S _d . The indicator S _W we consider proportional to the reactivity of the granules of the additive (with their equal porosity and degree of carbon graphitization).

Figure 5 shows the dependence of the specific surface of the granules on the diameter of the granules. From figure 5 it follows that a sharp rise in the indicator corresponds to the interval of the size of the granules 3-4 mm Curve 3 is the ratio of the Boudoir reaction time for granules of a given diameter to the residence time of the plate in the calculated volume of the furnace. Figure 6 shows the layout of four balls stacked in the correct chess packaging. Position 1 - balls of equal diameter, 2 - area of the through channel between the balls. The condition for the retention of the additive granules in a dry ore layer is the inequality d _gr <d _channel . Table 2 shows the calculation data for the diameter of the through channel using an approximate formula; d = 0,15d _{kan w.} This formula is suitable for the ore layer on the top, for which the diameter of the balls is taken equal to the average diameter of the balls throughout the volume of the layer.

Figure 7 shows the dependence of the diameter of the through channels on the diameter of the ball. To keep the granules of the additive in the ore layer not more than 0.5 cm, the diameter of equal balls does not exceed 40 mm (section for ore materials of Fig. 7). Section B with coarseness of equal balls d> 40 mm for dry coke provides the possibility of spilling granules of the additive with a diameter of not more than 0.5 cm through a layer of coke into the underlying layer of ore materials, which is necessary for the implementation of the method. According to the method for retaining the additive granules in the ore layer from the removal of blast furnace dust, the fraction of the additive granule fraction below 2 mm (the boundary for leaving the granules in the furnace) does not exceed 5% of the total dose of the additive. The influence of the shape factor of the pieces on the diameter of the through channels is not taken into account and, apparently, does not change the calculation results too much.

D. Conditions for the destruction of the shell of an isolated granule.

We assume that in the volume of the plate of FIG. 2, the volume of slag-forming materials is equal to 0.5 of the volume of the plate (there is no liquid phase). We believe that additives are evenly distributed inside the slag-forming granules (with a mixture uniformity coefficient of not more than two). For each granule, there is a volume of slag-forming, creating a shell of the granule, which has some tightness. When the Boudoir reaction proceeds, carbon monoxide is released from the granule, which creates an excess pore pressure of the gas inside the shell, which causes cracking of the shell. The intra-pore volume is attached to the initial porosity of the plate, reducing the gas pressure gradient along the height of the plate. A film of reduced iron is formed on the inner surface of the shell, which strengthens the shell framework (strength). To calculate the conditions for the destruction of the shell, we determine the volume of slag-forming materials per granule. The volume of slag-forming inside the plate is equal to V _sl = 0.6 × 0.5 = 0.3 m ³ / t h = 0.3 (10 ⁺⁶ ) cm ³ / t h. With an additive consumption of 100 kg / t h, the specific amount of slag per 1 kg of the additive will be:

V _beats = V _sl / S _add = 0.3 (10 ⁺⁶ ) / 100 = 0.3 (10 ⁺⁴ ) cm ³ / kg of the additive.

The diameter of the shell is calculated as follows. The volume of one granule of diameter d:

V _d = 0.52 d $\genfrac{}{}{0ex}{}{\text{3}}{\text{gr}}$ = 0.52 A (10 ^-3 ) cm ³ / gran, where A = (10 ⁺³ ) d $\genfrac{}{}{0ex}{}{\text{3}}{\text{gr}}$ - auxiliary parameter.

Weight of one granule of diameter d:

G _d = V _d γ _gp = 0.52 A (10 ^-3 ) × (10 ^-3 ) = 0.52 A (10 ^-6 ) kg / gran, where γ _gr = (10 ^-3 ) kg / cm ³ - the specific gravity of the substance of the granule. The number of granules of diameter d per 1 kg of additive:

N _d = 1 kg / G _d = 1 / 0.52 A (10 ^-6 ) = (10 ⁺⁶ ) 1.9 / A granules / kg of additive.

The volume of the shell per one granule of a given diameter

V _shell = V _beats / N _d = 0.3 (10 ⁺⁴ ) / (10 ⁺⁶ ) 1.9 / A = (10 ^-2 ) 0.16 / A cm ³ shell / gran.

Checking the dimension of the formula:

(cm ³ / kg add) / (gran / kg add) = (cm ³ / kg) × (kg / gran) = cm ³ envelope / gran.

Conventionally, we consider the shell volume equal to the volume of a continuous sphere with a diameter

where the first bracket after extracting the root of the third degree is 0.4 (10 ^-2 ) = 0.16. The second bracket after extracting the root of the third degree is

On Fig shows a curve of the diameter of the continuous shell from the diameter of one granule according to table 3. Figure 9 shows a diagram of the granule and the surrounding shell with a diameter of 3 mm

According to the method for destruction of the shell, the force of the interstitial pressure of the gas exceeds the strength of the shell material in a dangerous section. The degree of use of the additive substance for the Boudoir reaction depends on the diameter of the granule and the consumption of the additive. The required carbon consumption of the additive inside the plate does not exceed 50 kg / t h. The remaining carbon of the additive is spent on the Boudoir reaction in the dry zone, melting and drainage zones. We take the maximum degree of use of the additive substance in the plate for the Boudoir reaction η _pre = 0.5 / {S _add (10 ^-2 )} (curves 1 and 2 in figure 5). The numbers on the curves are the consumption of the additive kg / t h. The limiting degrees of use of the additive take into account the diameter of the granule, as well as the influence of the degree of graphitization of the carbon of the additive, porosity of the granules, and reaction temperature. Theoretical gas yield according to the Boudoir reaction per kg of additive V _theory = (22.4 / 12) 0.9 η _bud = 1.62 (10 ⁺⁶ ) η _bud cm ³ / kg, where 0.9 is the content of non-volatile carbon of the additive .

For one granule of a given diameter d at an additive flow rate of 100 kg / th, there is a gas output from the Boudoir reaction:

V _bud = V _theory G _gp = 1.62 (10 ⁺⁶ ) 0.52 A (10 ^-6 ) η _bud = 0.84 A η _bud

To determine the excess intrapore gas pressure, the Boyle-Mariotte formula is used, in which the initial volume and pressure of the gas are reduced to normal conditions (1 atm, 20 ° C). The process of gas accumulation in an isolated pore is considered isothermal.

Initial conditions: gas volume = V _bud , gas pressure in the volume V _bud is equal to P _bud = 1 atm.

Final conditions: the volume of compressed gas is equal to the volume of the pores after burning out the granules = V _d .

Excess pore gas pressure after the end of the Boudoir reaction = P _d .

We obtain P _d = P _bud (V _bud / V _d ), whence, substituting the initial data, we obtain:

P _d = 0.84 A η _Bud / 0.52 A (10 ^-3 ) = 1620 η _Bud atm / g.

Those. intrapore gas pressure depends on the diameter of the granule indirectly through the coefficient η _bud . During the accumulation of gas inside the pore, gas leaks through the porous shell from the slag-forming substances. Moreover, the degree of gas leakage decreases with increasing diameter of the granule and vice versa, due to the fall of η _bud (table 1), up to a pressureless pore volume with a granule diameter of 0.5 cm at the smallest shell thickness (table 3).

The leakage coefficient is determined by the formula η _ut = 1-0.5d _gp .

The practical coefficient of pressure reduction is η _pp = η _bud (1-η _ut ).

Table 4 shows the data of the inter-pore pressure of the gas depending on the diameter of the granule according to the above formula. We believe that the destruction of the shell occurs along the diametrical cross-sectional area - S _about , cm ² / about (Fig.9). The thickness of the hollow shell is conventionally equal to 0.5 of the diameter of the continuous shell of equal volume. Pneumatic forces applied to the horizontal projection of the granules S _gp are equal and opposite in sign to the _{tensile strength of the} shell along the diametrical area (Fig.9):

f _pnevm = P _bud S _g = 40.5 × 0.78 _{d gp} = 31.6 d _g kg / cm ² .

Table 5 shows the pneumatic forces of rupture of the shell depending on the diameter of the granules.

The strength of the shell is the product of the square shell at a diametral fracture stress sheath material - δ _discharge kg / cm ^2:

f _toughness = δ _bit S _r kg / rev

The tensile stress of the substance of the slag-forming substances is taken from experimental measurements of the tensile stress for clay rocks of different consistencies in the range of 50-500 kg / cm ² [9]. As a result of the Boudoir reaction, a thin film of spongy iron is created inside the shell, forming a reinforcing skeleton of the shell material and increasing its strength. Table 6 shows the data on the strength of the shell for the accepted values of tensile stresses taking into account the effect of a metallized skeleton on hardening of the shell. According to other fracture schemes, the shell strength will be lower than that accepted in this calculation. In contrast to the purely thermal destruction of ore-fuel granules [10], the Boudoir reaction in the plate creates pneumatic forces that mechanically destroy the shell. Figure 10 according to tables 5 and 6 are constructed curves of the balance of loads along the diametrical section of the shell. Curve 1 - pneumatic forces on the shell (table 5). Curves 2, 3, 4 - the strength of the shell along the diametrical section (the numbers on the curves are the tensile stresses of the shell material). From figure 10 it follows that when the diameter of the granules is higher than 0.2 cm, mechanical destruction of the shell along the diametrical section due to the intrapore gas pressure is ensured. The geometric pore volume from the burning of granules is:

Δε _add = (S _add η _bud ) / (V _pl γ _add ).

Substituting the initial data, we obtain the calculated formula for the increase in the porosity of the plate: Δε _add = (S _add η _Bud ) / (0.6 × 1000) = 0.0016 (S _add η _Bud ) m ³ / m ^{3 the} volume of the plate, where 1000 kg / m ³ is the specific gravity of the additive substance, 0.6 is the volume of the plate after shrinkage m ³ / t h, S _add is the consumption of the additive kg / t h. Table 7 shows the data on the geometric increase in the porosity of the plate after burning off the granules of the additive at the consumption of the additive in 50-100 kg / m h. The approximate relationship porosity ore on the throat plate and in the _count given by the formula ε = 2ε _mp. Figure 11 shows the dependence of the geometric increase in porosity after burning the granules of the additive on the diameter of the granules of the additive and the accepted maximum degree of use of the substance of the additive according to the Boudoir reaction. This dependence is necessary according to the method for balancing the available and necessary porosity in the plate, ensuring the movement of the column in the field of stability.

D. Calculation of gravitational forces applied from above to the plate.

The movement of the plate by the method is carried out due to the geometric increase in porosity from the burned-out granules, which reduces the gas pressure gradient along the height of the plate. Gravity forces from the dead weight of the plate are applied from above to the plate. Bottom is the gas pressure gradient along the height of the plate. Both values are assigned to one meter of the height of the calculated softening volume (Fig. 2) for commensurability with the gas pressure gradient at the top (first dimension) and to one cm ^{2 of the} hydraulic section of the furnace (second dimension). The gravitational force on its own weight is considered independent of the consumption of the additive due to the burning of granules when the plate moves down: f _pl = γ _pl N _pl K _A kg / cm ² m, where the density of the plate material γ _pl = (1500 / 0.6) (10 ^{- 4} ) = 0.25 kg / cm ² m; N _pl = 0.375 (Table 10); for the first variant of the method we obtain f _PL = 0.25 × 0.375 K _A = 0.094 K _A kg / cm ² m. Taking into account possible inaccuracies in the calculation, we roundly take the gravity force from the dead weight of the plate equal to (when working on one coke) f _PL = 0 , 01 kg / cm ² m. The transfer of forces from the overlying layers to the plate is not taken into account in this calculation. According to [11], with a coal dust consumption of up to 0.2 t / t h, pulsations of the temperature of the waste water from copper refrigerators installed in the walls of the lower part of the furnace were detected. These pulsations are a sign of intense gas flows creating a loose ring at the walls of the lower part of the furnace (item 9 in figure 2). The loose ring reduces friction against the walls, i.e. the coefficient of active weight of the plate increases. The effect of high coal dust consumption on pulsations is physically equivalent to the impact on pulsations of fine ore smelting. For these reasons, for both variants of the method, we take the coefficient of active weight of the plate K _A = 1. For the second variant of the method, the force of gravity of the plate is due to (at a constant porosity of the substance of the plate) an increase in the fraction of the height of the ore plate with respect to one meter of the calculated volume (Table 10).

E. Gas pressure gradient along the height of the plate (plug).

We determine the gradient of gas pressure in the plate with respect to the known gradient of gas pressure in the ore layer of the top. This ratio is inversely proportional to the ratio of the squared porosities of the substance in both layers and directly proportional to the ratio of their volume fractions in the calculated volume. For calculation: the heights of both calculated volumes are equal to one meter. Adopted unchanged: equivalent grain diameter, resistance coefficients of the substance, second gas flow rate in both layers. We obtain, under the indicated assumptions, the formula:

gR (P _pl / P _count ) = (0.3 / ε _pl ) ² (N _pl / N _count ) kg / cm ³ m.

G. Determination of the gas pressure gradient in the ore layer of the top. (Fuel - one coke). The calculation is carried out for a furnace with a volume of 2000 m ³ , a column height of -20 m, the total pressure drop of the gas (tuyeres / top) is 1.5 kg / cm ² . The average gas pressure gradient in the layer one meter high gRP _sl = 1.5 / 20 = 0.075 kg / cm ² m. We consider the pressure gradient approximately in the ore layer of the top of the furnace (when the volume fraction of ore in the top of the furnace top is _Ncol = 0.5 of the volume of the calculation layer) equal to gRP _sl = 0.05 kg / cm ² m.

The ratio of volume fractions of ore in the plate and on the top:

N _pl / N _count = 0.375 / 0.5 = 0.75.

Total: gR (P _PL ) = 0.05 × 0.75 (0.3 / ε _PL ) ² = 0.375 (0.3 / ε _PL ) ² kg / cm ² m.

Table 8 shows the data of the pressure gradient in the plate depending on the test porosity of the plate substance during the melting of small ore. It is conventionally accepted that the porosity in the ore layer of the top remains unchanged, the ore is crushed in the dry zone of the furnace, the ore shrinkage in the softening zone is 0.4.

H. The first variant of the method is the melting of small ore. Table 9 shows the data of the pressure gradient of the gas in the ore plate depending on the diameter of the granule, which creates a geometric increase in porosity (table 7) and is included in the total porosity of the plate according to the formula: ε _full = ε _test + Δε _add . The test porosity of the plate substance is taken as an indicator of the fraction of fine ore fractions and is determined from the Furnas curves. The relationship of the porosity of the ore on the top and in the plate ε _count = 2ε _prob . The geometric increase in porosity (table 7) is taken equal for all values of the test porosity for each diameter of the granule.

In Fig. 12, according to table 9, curves 1, 2, 3, 4 are plotted - the dependence of the gas pressure gradient in the plate on the test porosity of the plate material and the diameter of the granule (numbers on the curves). The practical minimum particle size of the granule is 2 mm. Line 4 corresponds to the pressure gradient of the gas in the plate with one coke fuel. Line 5 - gravity force from the dead weight of the plate (one coke). The intersection point of 6 lines 4 and 5 is practically reliable (the porosity of the substance of the plate is rounded 0.18). With such porosity (fuel - one coke), the load balance will be positive, and the column will move in a stable area. The permissible porosity of the plate according to the first variant of the method is 0.16 m ³ / m ³ at point 7. To maintain a positive load balance at this point, the diameter of the granules will be at least 3 mm. Moreover, the increase in available porosity from the burning of granules of this diameter should exceed the porosity deficit at this point equal to 0.02. According to Fig. 11, when the granule diameter is 2 mm, the available increase in porosity is 0.04 m ³ / m ³ , i.e. sufficient to maintain evenness. To the left of point 7, the deficiency of porosity of the plate is compensated by the method with an additive granule diameter of at least 2 mm. To the left of point 8, the porosity deficit is only partially compensated by the method, and it is supplemented by other effects on the movement of the column — the growth of the radius of the hole 2 in FIG. 2, the loading of larger coke into the hole 2 and the introduction of diluent additives into the ore part of the charge, which reduce the strength of slag-forming plates .

I. Second embodiment of the invention (coal dust plus additive). The operation of the furnace is disrupted when the consumption of coal dust is 0.2 t / t h due to several reasons: underburning of coal dust in the volume of the circulation cavity, a sharp increase in the fraction of the plate in the estimated softening volume (Fig. 2), which is confirmed by the growth of the gas pressure gradient along the height of the softening zone [12, 13]. In Fig.13 shows the growth curve of the ore load on the residual coke minus coke filling the hole 2 in Fig.2. The kink of the curve corresponds to the level of consumption of coal dust 0.2-0.3 t / t h.

Table 10 shows the balances of the gravity of the plate depending on the consumption of coal dust:

f _PL = γ _PL N _PL K _A , where γ _PL = 0.25 kg / cm ² m is the density of the plate substance during shrinkage of ore in the plate, equal to 0.4.

On Fig shows the relationship of the density and porosity of the substance of the plate during shrinkage of ore in the range from zero to 0.4.

Table 11 shows the pressure gradients of the gas in the ore layer of the top depending on the consumption of coal dust gRP _count = 0.1 N _count kg / cm ² m

Table 12 shows the gas pressure gradients along the height of the plate according to the formula:

gRP _pl = gRP _count (N _pl / N _count ) (0.3 / ε _count ) ² kg / cm ² m.

Table 13 shows the coefficient of gravitational load on the plate: K _PL = f _PL / gRP _PL depending on the test porosity of the plate (without additives). On Fig shows the dependence of K _PL from the consumption of coal dust. Line 1 corresponds to a theoretically neutral balance of plate loads. Lines 2, 3, 4 - values of the coefficient K _PL with different test porosity of the substance of the plate (without additives). Point 5 of the intersection of lines 1 and 3 is a reliable value of the consumption of coal dust of about 0.2 t / t h, corresponding to the test porosity of the plate 0.185 m ³ / m ³ . Due to inaccuracies in the calculation, we roundly accept porosity at a point equal to 0.18 m ³ / m ³ . To the left of point 5, the load balance is positive, to the right - negative, which coincides with the abrupt violations of the furnace course that were practically detected when coal dust consumption was over 0.2 t / t h [12, 13]. The rounded porosity of 0.18 is close to the porosity of the plate when working on the same coke (Fig). When the load factor of the plate K _PL = 1 and the evenness of the furnace without entering additives is unattainable. To ensure evenness in the range of coal dust consumption of 0.2-0.4 t / t h, according to the method, an additive is introduced into the ore with a pellet diameter of not more than 4 mm, at which the available increase in porosity of the plate exceeds the porosity deficit (Fig. 11). In this case, the load factor will be positive in the indicated interval of the consumption of coal dust K _PL = 1.

On Fig shows the relationship of the available increase in porosity from the introduction of additives and deficiency of porosity for two variants of the method. The accepted granule diameter is 3 mm, average for a range of granule diameters from zero to 5 mm. In Fig. 16, line 1 corresponds to the increase in disposable porosity when smelting a fine charge for a granule diameter of 3 mm (the first variant of the method). Permissible test porosity of the plate with this diameter of the granule will be less than 0.16 m ³ / m ³ . If the porosity of the plate is below 0.16, the available porosity is not enough to compensate for the deficiency of porosity and the course of the furnace will be unstable. Line 2 corresponds to the increase in available porosity at a consumption of coal dust of up to 0.4 t / t h and a porosity deficit of at least 0.01 m ³ / m ³ (Fig. 15). With such a consumption of coal dust and a decrease in the test porosity of the plate below 0.18, the method only partially compensates for the deficiency of porosity of the plate and other effects on the movement of the column should be used. Line 3 corresponds to the joint effect on the porosity deficit of fine ore smelting and the introduction of coal dust with a flow rate above 0.2 t / th. In this case, the introduction of granules with an average diameter of <3 mm will become insufficient to compensate for the porosity deficit, and such a melting mode is not feasible.

K. Curve of thermal stability of melting.

The accepted stresses of the rupture of the shell material (table 6) are due to the preservation of the constant enthalpy of slag-forming. On Fig shows a curve of thermal stability of the heat in graphical form ROSA (by the type of wind rose diagrams). The abscissa _{shows the} Si content of _iron . On the ordinate axis is the thermal value of the top gas equal to the enthalpy of gas left by it in the furnace. Both parameters are observable and can serve for monitoring purposes. Circle 1 - the permissible limits of the process. Circle 2 is the boundary of stability. Inside circle 2, the thermal capacity of the furnace is sufficient to compensate for fluctuations in both parameters without external influences on the process. In modern practice, the most dangerous are the cooling of the furnace. Point 3 shows the start of a cold snap. Line 4 - the development of cooling, in which Si _{cast iron} and the thermal value of the gas rapidly fall. The circular cycle 5 corresponds to impaired column motion (cliffs). In this case, the thermal capacity of the furnace will not be enough to compensate for the process. If point 3 is missed, hidden from visual control, the process develops quickly and cannot be synchronously compensated from above. And from below, there may be no possibility of increasing enthalpy of blast. In this case, a decrease in ore load, blast consumption, etc. is inevitable. With temporary disturbances, these effects are canceled after the process returns to area 2. With constant violations, the impacts are maintained at a new level. According to the method, a computer program for constructing the curve of Fig. 18 and current monitoring of the position of the process operating point in region 2 by the method [5] are implemented. In this case, the tensile stresses of the shell substance correspond to the values accepted in the calculation (table 6).

Examples of the method.

The first example.

The implementation of the method is described in relation to a furnace with a volume of 2000 m ³ with the above parameters of the charge and the melting mode. The furnace is equipped with a computer program for calculating the evenness coefficient (Fig. 3A) and thermal stability control (Fig. 17), a device for introducing a carbon-containing additive with a particle size of 2-5 mm into the loading path to the top (Fig. 1). The top diameter of this furnace is 7.3 m, the working height is 20 m. The position of the upper boundary of the calculated volume is 4.7 m above the tuyere level. We believe that the height positions of the centers of gravity of the real and estimated softening volumes coincide. Coke - standard, porosity at the top of 0.5 m ³ / m ³ . An indicator of the quality of ore materials is their porosity on the top. The relationship of porosity at the top and the fraction of fine fractions in ore is determined by the Furnas curve [15]. When the porosity of ore materials is below the norm of 0.3 m ³ / m ^{3, the} porosity of the plate also decreases. The temperature of the onset of softening of the ore is in the range of 800-1000 ° C and coincides with the position of the gas isotherm at 1000 ° C at a level of 4.7 m above the level of the tuyere. The radius of the Central outlet 2 (figure 2) is 1.4 m. Cast iron - conversion composition (percent): Si _{cast iron} = 0.5, Mn = 0.2, P = 0.1, S = 0.02. The coefficient of evenness ρ _x = 0.5 (line 1, figure 3). According to the method, a carbon-containing additive (fractions of 2-5 mm) is introduced into each component of the ore part of the charge before loading on the top with a total consumption of 10 kg / t of additive for every 0.01 decrease in the coefficient of evenness below the permissible level of 0.5. Deviations of Si _{cast iron} do not exceed 0.05% (10% of the nominal value inside circle 1 of FIG. 17). The porosity of the plate when working on one coke is 0.18 m ³ / m ³ (point 6, Fig. 12). The decrease in the test porosity of the plate to 0.16 m ³ / m ³ corresponds to the melting of small ore. According to the method, when the size of the additive is 3 mm or lower, the increase in disposable porosity is 0.02 m ³ / m ³ , which is close to the porosity (0.18-0.16) = 0.02 m ³ / m ³ , necessary to create a positive balance gravitational forces on the plate (line 1, Fig.16). In this case, the coefficient of evenness will be ρ _x ≤ 0.5 (line 3, Fig. 3). When reducing the test porosity of the plate below 0.16 m ³ / m ³ (Fig. 12), the method does not fully ensure the stability of the melt and other effects on the process are necessary (while maintaining the mode of the method): increasing the radius of the hole 2 in Fig. 2, loading larger coke , the creation of a loose ring 9 on the furnace top and others (figure 2).

An example of the second.

The volume of the furnace and the melting mode are the above. Coal dust consumption in the range of 0.1-0.4 t / t h. Additive consumption in the range of 50-100 kg / t h, over the total consumption of coke and coal dust. The furnace is equipped with a computer program for calculating the coefficient of evenness and thermal stability of the process, a device for inputting coal dust in a blast with a capacity of 400 t / h, finely ground dust of 50 microns, ash ash of not more than 20%, a device for inputting a carbon-containing additive into the loading path (Fig. 1) . Figure 2 shows the lower boundaries of a compact ore plate in a calculated volume of one meter high depending on the consumption of coal dust (figures on the curves). The residence time of the plate in the softening volume is in the range of 30-60 minutes, which overlaps the time for the complete completion of the Boudoir reaction inside the plate. The Si content of the _{cast iron is} kept stable (FIG. 17). To control the level of consumption of additives, we determine the coefficient of evenness, which is above the permissible level (line 3, figure 3). The normal porosity of the plate, ensuring evenness, is 0.18 m ³ / m ³ . With an increase in the consumption of coal dust to 0.4 t / t h, the porosity deficit of the plate is 0.02 m ³ / m ³ (Fig. 16, line 2). In Fig. 16, line 2 shows the available increase in porosity of the plate when introducing a carbon-containing additive and the average particle size of the granules of the additive is 3 mm (Fig. 11), equal to 0.02 m ³ / m ³ , which is close to the porosity deficit at a coal dust consumption of 0.4 t / t h. According to the method, spikes of ore and residual coke are collected in compact layers within 1-5 spikes in each compact. The number of ears in the compact N _comp = 0.5 / V _oc , where V _oc is the residual volume of coke, on the top m ³ / t h, N _comp is the number of residual coke layers in the compact. With a residual coke consumption of 0.1 t / t h N _comp = 5. The ore compact volume is equal to the spike volume multiplied by N _comp . In figure 2, lines 6, 7, 8 correspond to the boundaries of the compacts in the calculated softening volume (the numbers on the curves indicate the consumption of coal dust). With the combination of increased consumption of coal dust in excess of 0.2 t / t h, fine ore, the porosity deficit in Fig. 16 (line 3) is 0.04 m ³ / m ³ . The method compensates for 0.02 m ³ / m ³ deficiency of porosity. To maintain the stability of the melt in this case, they influence the loading mode on the top (increasing the radius of the hole 2 in FIG. 2), creating an additional loose peripheral ring. On Fig shows the change in the porosity of the charge components along the height of the column (arbitrary scale): 1 - dry zone, 2 - estimated softening volume 1 m high, 3 - melting volume, 4 - drainage zone, 5 - coke porosity in the dry zone, 6 - the porosity of the layer of irrigated coke in the drainage volume, 7 - the porosity of the ore in the dry zone, 8 - the test porosity of the plate 0.18 m ³ / m ³ with one coke or at a consumption of coal dust to the level of 0.2 t / t h 9-10 - abnormal porosity of the plate with fine ore, 11 - porosity of the plate by the method, 12 - porosity of the melt in the drainage zone. The accepted equivalent fuel consumption of 0.5 t / t h is reduced when a carbon-containing additive is introduced by replacing part of the coke consumed by the Boudouard reaction in the plate up to 0.05 t / t h from the accepted initial level.

LITERATURE

1. Kakubo et al. Investigation of the domain process with mixed loading of charge materials: Express information Central Research Institute of Information and feasibility studies of ferrous metallurgy. M., 1984, issue 21, p. 1-3. Testo to Hagone 1984, vol. 70, N4, p. 84-550 (Japan).

2. RU 2042714 C1, C 21 B 5/00, 08.27.1995.

3. Bazilevich S.V. and other Agglomeration. - M.: Metallurgy, 1967, p. 311.

4. Pokhvisnev A.N. and others. On the minimum consumption of coke in a blast furnace under the conditions of the gas dynamics of the process. Steel, 1969, N12, p. 59-61.

5. Rakovsky B.M. etc. Blast furnace melting modes under unstable working conditions. Ferrous metallurgy. Bulletin of scientific, technical and economic information. M., 1999, N9-10, p. 5-20; N1l-12, p. 19-28.

6. Efimenko G.G. Determination of the composition of primary slag in a blast furnace. Iron metallurgy. Scientific works of DMI. Kharkov, Moscow, issue 29, p. 256-274.

7. Volovik G.A. and others. On the issue of softening temperatures of ores and agglomerates. Iron metallurgy. Scientific works of DMI. Kharkov, Moscow, issue 29, p.105-134.

8. Lyuban A.P. Analysis of the phenomena of the domain process. - M.: Metallurgizdat, 1955, p. 191-192.

9. Churinov M.V. Handbook of engineering geology. - M .: Nedra, 1968, p. 69, tab. 33.

10. Hodak L.Z. et al. Changes in the phase composition and the mechanism of primary slag formation during smelting of fuel-oil granules. Proceedings of the Institute of Combustible Minerals. - M .: Akademizdat, 1963, v.22, p. 79-92.

11. Helenbrook R. Requirements for the cooling of copper blast furnace refrigerators. Helenbrook R.G. et.al. Water requirements for blast furnace copper staves. Jron and Steelmaker, 2000, N6, p. 45-51.

12. Okochi et al. Achievement of coal dust consumption of 266 kg / t h on a blast furnace N3 of the Fukuyama plant. Ososhi et.al. Achievement of high rate pulverized coal injection of 266 kg at Fukuyama N3 B.F. 4th European Coke and Ironmaking congress. Paris, 2000, pp. 198-202.

13. Nozawa K et al. Operational practice of the N1 blast furnace at Kakogawa plant with a consumption of coal dust of 250 kg / t. Nozawa K. et.al. Practical operations at an ultra high coal injection rate over 250 kg / t HM in Kokagawa N Blast Furnace. International conference on new development in metallurgical process. Technology Proceedings. Dusseldorf, 1999, p. 87-90.

14. Taylor A.G. et al. Expert assessment method for predicting the unstable operation of a blast furnace at British STIL plants. Taylor A.G. et.al. Advanced signal processing methods and expert system development for predicting and assessing blast furnace instability within British Steel. International conference on new development in metallurgical process. Technology Proceedings. Dusseldorf, 1999, p. 98-102.

15. Furnas C.C. The movement of gases through a layer of bulk materials. DOMEZ. Kharkov - Dnepropetrovsk, 1932, p. 74-87.

Claims (5)

1. The method of blast furnace smelting, comprising separate loading through the top of iron ore materials and coke, loading of a carbon-containing additive, blowing blast and coal dust into the tuyeres, characterized in that before loading onto the furnace top, iron ore materials are mixed with a carbon-containing additive in an amount of 10-100 kg / t cast iron in excess of the total fuel consumption, loaded through the top and blown into the tuyeres, while the degree of homogeneity of the mixture of iron ore materials and carbon-containing additives is not more than two, and the size of carbon containing additives does not exceed 5 mm. 2. The method according to claim 1, characterized in that coke waste, crushed anthracite, granule, a water-coal suspension are used as a carbon-containing additive. 3. The method according to claim 1, characterized in that the average particle size of the iron ore materials is not more than 40 mm, the size of the coke is not less than 40 mm 4. The method according to claim 3, characterized in that they determine the coefficient of evenness of the furnace and when the porosity of iron ore materials is below 0.3 m ³ / m ^{3 the} consumption of carbon-containing additives is changed in doses of 10 kg / t of pig iron for every 0.01 deviation of the coefficient of evenness below the acceptable level of 0.5. 5. The method according to claim 1, characterized in that when the flow rate of coal dust blown into the tuyeres is in the range of 0.2-0.4 t / t of pig iron, the consumption of carbon-containing additives is 50-100 kg / t of cast iron, while the coke supply and iron ore materials are loaded through the top with compact layers within the range of 1-5 heads in each layer.

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