A method for obtaining the initial support force of the advance unit support in mining tunnel

Abstract

The present invention discloses a method for obtaining the initial support force of an advance unit support of a mining tunnel, and belongs to the field of advance unit supports of mining tunnels. Specifically, it includes obtaining parameters of the mining tunnel and the advance unit support; determining the initial support force of the advance unit support under load; determining the initial support force of the advance unit support in a steady state; determining the initial support force of the advance unit support under destruction; and determining a reasonable initial support force of the advance unit support through the initial support force of the advance unit support under load, the initial support force of the advance unit support in a steady state, and the initial support force of the advance unit support under destruction. The present invention fully considers the three on-site work dimensions of the load-bearing requirements, position stability, and avoiding the destruction of active support of the advance unit support, and forms a multi-index composite comparison system for solving the reasonable initial support force of the advance unit support of the tunnel. This method not only considers the load that the initial support force of the advance unit support needs to bear in the initial support stage, but also fully considers the position of the advance unit support itself and its influence on the active support, which is more in line with the actual needs on site.

Description

Technical Field

[0001] The invention relates to the field of an advance unit support for a mining tunnel, and in particular to a method for obtaining the initial support force of an advance unit support for a mining tunnel. Background Art

[0002] As the throat of safe production of working face, the advance support of mining tunnel has gone through the support process of wooden support, single hydraulic support, chute hydraulic support and advance unit support.

[0003] In the past, advance support for mining tunnels was mostly based on "anchor cable + chute support group" as the advance support method; that is, on the basis of active support by anchor cables, chute supports were used to control surrounding rock deformation. However, for chute supports, since the chute supports need to be moved by the hydraulic jacks between the frame groups, the supports will inevitably support the roof repeatedly. In addition, the top beam area of ​​the chute support is large and the working resistance of the frame group is high. The top beam will bend, fold, and damage the exposed head of the anchor cable, which will lead to damage to the anchor cable support system, seriously affecting the quality of the roadway roof support, and even causing roof accidents and casualties.

[0004] The advanced unit support can effectively avoid the damage to the roof caused by repeated support due to its flexible frame moving measures. In addition, the top beam area of ​​the advanced unit support is small and the layout method is flexible, and the top beam has little effect on the anchor cable support system. Therefore, if the hydraulic support support of the drift is still used, it will not only have an adverse effect on the stable support of the surrounding rock of the advanced section of the tunnel, but also affect the recovery efficiency of the working face to a certain extent. The use of advanced unit supports can solve this problem well, but through a large number of literature searches, it is found that there is no theoretical research and practical report on the reasonable initial support force design of the advanced unit support. At present, the design of the initial support force of the advanced unit support is still in a blank stage. Therefore, it is urgent to propose a theoretical research method for the design of the initial support force of the advanced unit support. Summary of the invention

[0005] In order to solve the problem of accurately obtaining the initial support force of an advance support of a deep-buried mining tunnel, the present invention provides a method for obtaining the initial support force of an advance unit support of a mining tunnel.

[0006] The present invention adopts the following technical solutions:

[0007] A method for obtaining the initial support force of an advance unit support in a mining tunnel comprises the following steps:

[0008] Step 1: Obtain the parameters of the mining tunnel and the advance unit support.

[0009] Step 2: Using the parameters of the mining tunnel and the advance unit support, determine the initial support force F of the advance unit support P1。

[0010] Step 3: Determine the steady-state initial support force F of the leading unit support P2。

[0011] Step 4: Determine the initial support force F of the leading unit support P3。

[0012] Step 5: Carry the initial support force F through the advance unit support P1 , Steady-state initial support force F of the leading unit support P2 and the initial support force F of the leading unit support P3 , determine the reasonable initial support force F of the leading unit support P .

[0013] Preferably, the parameters of the mining tunnel and the advance unit support include the elastic-plastic boundary length of the tunnel side, the uniformly distributed load above the rock beam, the length of the rock beam, the bulk density of the overlying rock strata, the distance between the tunnel advance unit support and the coal wall, the ultimate yield strength of the anchor rod, the tensile strength of the roof rock beam and the bending section modulus of the rock beam.

[0014] The method for obtaining the length of the elastic-plastic boundary of the tunnel side is as follows: at the leading section of the working face of 150m, a double-end water plug is placed in the middle of the two sides of the tunnel by horizontal drilling. As the mining progresses, at a distance of 120m from the double-end water plug, a deep measurement is first made at the borehole at an equal distance of 5m, and the water injection loss at different hole depth sections is compared to preliminarily observe the crushing of the coal bodies on the two sides. When the ratio of the front and rear leakage is not less than 1.5, the exploration distance is reduced within this section, and the exhaustive method is used to determine that the difference between the current and rear values ​​is within 1m. When the ratio of the front and rear leakage is still not less than 1.5, it is considered that the junction of the front and rear sections is the elastic-plastic boundary of the two sides of the tunnel, that is:

[0015] (1);

[0016] (2);

[0017] Where n is the number of injection sections, Q is the amount of injection loss, and m is 3 ;x i x is the distance of each exploration hole of the double-ended water plug, m; 0 is the elastic-plastic boundary length of the tunnel side, m.

[0018] It should be noted that:

[0019] (1) Different from the double-ended water plug used in measuring the water-conducting fracture zone of the roof, the double-ended water plug used in measuring the side of the roadway in the present invention has a smaller borehole diameter than the roof borehole, generally 60-80 mm, to reduce damage to the coal body of the roadway side. In addition, since the degree of coal fragmentation of the roadway side is different, a reinforcement baffle is promptly arranged after drilling.

[0020] (2) To ensure the accuracy of the measurement results, after every three measurements, the instrument is removed from the drill hole and the integrity of the rubber occluder is inflated (or filled with water) to prevent data distortion caused by the instrument itself.

[0021] Preferably, the leading unit support bears the initial support force F P1 The calculation formula is:

[0022] (3);

[0023] In the formula, is the uniformly distributed load above the rock beam; is the length of the rock beam; is the tensile strength of the roof rock beam; is the bending section modulus of the rock beam; b is the distance between the tunnel advance unit support and the coal wall, specifically, b is the smaller value of the distance between the advance unit support and the coal walls on both sides.

[0024] In some tunnels, the tensile strength of the roof rock beam, the bending section modulus of the rock beam and the uniformly distributed load above the rock beam are difficult to accurately obtain, and some parameters are difficult to measure; considering that the parent rock layer above the two walls interacts with the potential separation layer, the weight of the potential separation layer is borne by the unit support and the parent rock layer. Considering the dangerous limit state, the weight of the potential separation layer is borne entirely by the unit support. By substituting the tunnel parameters under general conditions for calculation, the initial bearing force obtained is close to and slightly less than the weight of the anchoring layer coal rock mass within the roof range above the tunnel, that is, the expression is as follows:

[0025] ;

[0026] Where, B is the tunnel width, m; D is the row spacing of the advance unit support; L′ is the thickness of the anchoring rock layer, m; γ is the bulk density of the overlying rock layer, kN / m 3 .

[0027] Preferably, step 3 specifically includes the following steps:

[0028] Step 3.1: Obtain the horizontal force transmitted by the two sides of the roadway to the leading unit support , and its calculation formula is as follows:

[0029] F x = AS [ e − 2 x 0 ' tg φ mA ( K γ H + C cot φ ) + C cot φ ] / 100 (4);

[0030] in, ; A is the pressure coefficient, , μ is Poisson's ratio; S is the size of the base of the leading unit support, m 2 ; K is the stress concentration coefficient; H is the depth of coal seam; γ is the bulk density of the overlying rock layer, kN / m 3 ; is the weighted internal friction angle between the direct top and the direct bottom of the roadway; C is the weighted cohesion between the direct top and the direct bottom of the roadway; m is the average thickness of the coal seam; x 实体煤 is the elastic-plastic boundary length of the tunnel side close to the solid coal; x 运输 It is the elastic-plastic boundary length of the tunnel side near the belt conveyor.

[0031] Step 3.2: According to the principle of static balance, the advance unit support can only ensure its own position stability when the horizontal friction force on the advance unit support and the horizontal force generated by the deformation of the roadway coal wall are balanced. That is, the initial support force provided by the advance unit support at this time is the steady-state initial support force. The steady-state initial support force F of the advance unit support is determined according to the following formula: P2 :

[0032] (5);

[0033] In the formula, is the friction force between the top beam of the advance unit support and the top plate of the tunnel; is the friction coefficient of the contact surface between the advance unit bracket and the top plate.

[0034] Preferably, according to the stress state of the anchor rod and anchor cable by the top beam of the advance unit support, at the beginning, the anchor rod is mainly subjected to compressive stress; as the advance unit support continues to apply force, the anchor rod bends until it is embedded in the top rock layer, and it is mainly subjected to compressive stress at this stage; when the force applied by the advance unit support to the anchor rod is greater than the tensile strength of the support rod body, the anchor rod fails. The initial support force F of the advance unit support is determined according to the following formula P3 ;

[0035] (5);

[0036] In the formula, F P3 is the pressure applied by the advance unit support to the anchor rod, i.e., the initial support force of the advance unit support. According to the site conditions, the advance unit support can apply force to two anchor rods at most at the same time; S 杆体 is the nominal cross-sectional area of ​​the anchor rod body; σ s is the ultimate yield strength of the anchor rod.

[0037] Preferably, step 5 specifically includes the following steps:

[0038] Step 5.1: Determine the roof equivalent parameter [BQ] and the coal body equivalent parameter σ.

[0039] The roof equivalent parameter [BQ] is a comprehensive evaluation of the rock mass quality of underground engineering projects, and its expression is as follows:

[0040] (7);

[0041] [ BQ ] = BQ − 100 ( K 1 + K 2 + K 3 ) (8);

[0042] Where, BQ is the basic rock mass quality evaluation index; R C is the saturated uniaxial compressive strength of rock; K V is the rock mass integrity index; K 1 K is the correction factor for groundwater impact on underground projects; 2 K is the correction coefficient for the influence of the main structural surface of the underground project; 3 is the correction factor for the initial stress state.

[0043] During the calculation of equations (7) and (8), the following provisions should be met:

[0044] 1) When R C >90K V +30, R C =90K V +30 and K V Substitute and calculate the BQ value.

[0045] 2) When K V >0.04R C +0.4, K V =0.04R C +0.4 and R C Substitute and calculate the BQ value.

[0046] The coal mass equivalent parameter σ is a comprehensive evaluation of the mechanical properties of the rock mass, and its expression is as follows:

[0047] (9);

[0048] in, , , σ c is the uniaxial compressive strength of coal; GSI is the coefficient of coal fracture development; d is the mining impact index; is a media constant; is the rock hardness coefficient, is the rock crushing coefficient.

[0049] Step 5.2: Determine the weight coefficient. Determine F based on the roof equivalent parameter [BQ] and the coal body equivalent parameter σ P1 and F P2 The corresponding weight, a1 represents F P1 The weight of b1 represents F P2 The specific weights are shown in Table 1.

[0050] Table 1 Weight impact factor distribution table

[0051]

[0052] Step 5.3: Calculate the initial support force F of the weighted advance bracket P ’ .

[0053] The weight factors a1 and b1 form a matrix A;

[0054] A = [ a 1 b 1 ] ;

[0055] The determined leading unit support is subjected to the initial support force F P1 and the steady initial support force F of the leading unit support P2 Composition matrix B;

[0056] ;

[0057] Calculate the initial support force F of the weighted advance unit support P ’ ; (10).

[0058] Step 5.4: Compare the weighted leading unit support initial support force F P ’ and the initial support force F of the leading unit support P3 , taking full account of the safety of coal mines, the maximum value is taken as the reasonable initial support force F of the leading unit support P ;

[0059] (11).

[0060] The present invention has the following beneficial effects:

[0061] (1) The present invention proposes for the first time a method and idea for calculating the initial support force of the advance unit support, which realizes the scientific and quantitative design of the initial support force of the advance unit support under different types of tunnel conditions, and is beneficial to the stability control of the surrounding rock of the advance tunnel.

[0062] (2) The initial support force calculation method mentioned in the present invention fully considers the three on-site work dimensions of the advance unit support bearing requirements, position stability and avoiding damage to active support, forming a multi-index composite comparison system for solving the reasonable initial support force of the tunnel advance unit support. This method not only considers the load that the initial support force of the advance unit support needs to bear in the initial support stage, but also fully considers the position of the advance unit support itself and its impact on active support, which is more in line with the actual needs on site.

[0063] (3) The present invention innovatively combines the design of the initial support force of the advance unit support with the geological conditions of the tunnel, which is beneficial to the targeted initial support force design of the advance unit support for tunnels with different geological conditions. It has high practicality and universality, and can significantly improve the support level of the advance section tunnel.

[0064] (4) The calculation method mentioned in the present invention is simple, easy to use on site, and has good use effect. BRIEF DESCRIPTION OF THE DRAWINGS

[0065] Figure 1 This is the layout diagram of the elastic-plastic boundary monitoring points on both sides of the tunnel.

[0066] Figure 2 for Figure 1 Layout diagram of the medium AA monitoring section.

[0067] Figure 3 This is the plan layout of the 2301 track chute in Example 1.

[0068] Figure 4 This is the support cross-section diagram.

[0069] Figure 5 This is the curve of water loss from drilling holes on the side of the tunnel.

[0070] Figure 6 Schematic diagram of the double-ended water shutoff device used in on-site measurements.

[0071] Figure 7 This is a schematic diagram of the planed surface of the lane side when the double-ended water shutoff device is in use.

[0072] In the figure, 1. mining tunnel; 2. goaf; 3. double-end water plug; 4. advance unit support; 5. belt conveyor; 6. direct top; 7. direct bottom; 8. anchor rod; 9. water injection pipe connector; 10. rubber plug; 11. water injection probe; 12. water outlet; 13. reinforcement baffle; 14. water injection pipe; 15. inflation pipe. DETAILED DESCRIPTION

[0073] On the basis of fully considering the deficiencies of previous theoretical research and engineering practice, the present invention proposes a multi-index comprehensive determination method for the initial support force design of the tunnel advance unit support based on the physical and mechanical parameters of the tunnel surrounding rock, the elastoplastic boundaries of the two sides of the tunnel and the ultimate tensile strength of the anchor rod. Firstly, the physical and mechanical parameters of the surrounding rock are used as evaluation indicators to determine the basic conditions of the tunnel where the advance unit support is arranged, and the initial support force required to be borne by the advance unit support is calculated accordingly; secondly, during the arrangement of the advance unit support, when it is arranged in a mining tunnel with low surrounding rock strength, as the working face is mined, the stress of the overlying rock strata is redistributed, thereby forming lateral support pressure on the side of the tunnel, causing deformation of the surrounding rock; although the sinking of the tunnel roof and the bulging of the bottom plate will also affect the position of the advance unit support, in the process of the initial support force of the advance unit support, the influence of the deformation of the two sides on it is more obvious, so the solution of the horizontal force generated by the deformation of the two sides of the tunnel is the most intuitive and fundamental prerequisite for determining that the advance unit support maintains a stable position in the initial support stage, , the present invention uses the elastic-plastic boundary of the two sides of the tunnel as the second evaluation index to determine the initial support force required for the advance unit support to ensure steady state; thirdly, when the anchor support is too dense or the spacing between the advance unit supports is unreasonably designed, the advance unit support may destroy the original anchor support of the tunnel roof, causing the anchor failure. The initial support force of the advance unit support to destroy the anchor support is determined by solving the force required for the anchor to reach the ultimate tensile strength; finally, the physical and mechanical parameters of the tunnel surrounding rock are the basic mechanical parameters for calculating the bearing weight required for the initial support force of the advance unit support. The elastic-plastic boundary of the two sides is the key factor in defining whether the advance unit support can maintain position stability, and the ultimate tensile strength of the anchor is the direct cause of resisting the damage caused by the unreasonable support of the advance unit support. It is an important guarantee to ensure that the initial support force of the advance unit support does not destroy the active support. However, under different geological conditions, the contribution rate of the three indicators to the most reasonable initial support force of the advance unit support is different. Based on this, a composite evaluation standard under different geological conditions is designed to comprehensively determine the reasonable initial support force of the advance unit support. Therefore, the physical and mechanical parameters of the tunnel surrounding rock, the elastic-plastic boundaries of the two sides of the tunnel and the ultimate tensile strength of the anchor are used as evaluation indicators to determine the most reasonable initial support force of the advance unit support, which can not only ensure that the advance unit support controls the weight of the roof rock layer but also maintain the stability of the support's own position without destroying the active support structure, filling the gap in the method of determining the initial support force of the tunnel advance unit support.

[0074] The specific implementation of the present invention is further described below in conjunction with the accompanying drawings and specific embodiments: Example

[0075] The 2# coal seam is mainly mined at the 2301 working face of a certain mine. The coal seam is buried at a depth of H 808m, with an average coal seam thickness of m of 6.8m. The immediate roof 6 is siltstone with an average thickness of 6.3m, and the immediate bottom 7 is mudstone with an average thickness of 1.29m. The width of the mining tunnel B is 5.4m, the spacing D of the unit support layout is 6m in both tunnels, the thickness of the anchor layer L′ is 2.65m in both tunnels, and the bulk density of the overlying rock stratum γ is 13kN / m 3 The length of the rock beam l is 6m.

[0076] A double-ended water plug 3 was installed by drilling holes 150 m ahead of the mining tunnel 1 of the 2301 working face. When mining was 30 m, the elastic-plastic boundary length x of the tunnel side of the mining tunnel 1 close to the solid coal at the front 120 m was measured. 实体煤 is 23.6 m, and the elastic-plastic boundary length x of the tunnel side near the belt conveyor 5 is 运输 It is 36.8m.

[0077] A method for obtaining the initial support force of an advance unit support in a mining tunnel comprises the following steps:

[0078] Step 1: Obtain the parameters of the mining tunnel 1 and the advanced unit support 4. In this embodiment, the elastic-plastic boundary length of the tunnel side is 36.8m, the uniformly distributed load above the rock beam is 1230kN / m, the length of the rock beam is 6m, and the bulk density of the overlying rock stratum is 13kN / m 3 , the distance between the tunnel advance unit support and the coal wall is 2m, the tensile strength of the roof rock beam is 3MPa and the bending section modulus of the rock beam is 1.4m 3 .

[0079] The structure of the double-ended water plug is the prior art. Figure 6 and Figure 7 shown.

[0080] Step 2: Using the parameters of the mining tunnel and the advance unit support, determine the initial support force F of the advance unit support P1 .

[0081] In this embodiment, the leading unit bracket bears the initial support force F P1 for:

[0082] .

[0083] During the actual on-site investigation, if the basic physical parameters such as the rock properties, cross-sectional shape, and moment of inertia of the mine roof are difficult to determine, it is difficult to determine the tensile strength σ of the roof rock beam. t , rock beam bending section modulus W Z The uniformly distributed load q on the rock beam is accurately determined. Then the initial support force F of the leading unit support is P1 The approximate solution is:

[0084] .

[0085] Step 3: Determine the steady-state initial support force F of the leading unit support P2 .

[0086] Step 3.1: In this example, A is 0.43, K is 2, and S is 1.70m. 2 , C is 1.80MPa, ψ is 32.08°, then the horizontal force F transmitted by the two sides of the tunnel to the advance unit support X for:

[0087] F x = AS [ e − 2 x 0 tg φ mA ( K γ H + C cot φ ) + C cot φ ] / 100 = 0 . 43 × 1 . 70 × [ e − 2 × 36 . 8 × tg 32 . 08 ∘ 6 . 8 × 0 . 43 ( 2 × 13 × 808 + 1800 × cot 32 . 08 ∘ ) + 1800 × cot 32 . 08 ∘ ] / 100 = 21 . 0 kN .

[0088] Step 3.2: Based on the principle of static balance, Take 0.85, and then get the steady-state initial support force F of the leading unit support P2 for:

[0089] .

[0090] Step 4: Determine the initial support force F of the leading unit support P3 .

[0091] The ultimate yield strength of the anchor rod 8 used is KPa, nominal cross-sectional area S corresponding to anchor rod 8 杆体 3.8×10 -4 m 2 , then the initial support force F of the leading unit support is P3 for:

[0092] .

[0093] Step 5: Carry the initial support force F through the advance unit support P1 , Steady-state initial support force F of the leading unit support P2 and the initial support force F of the leading unit support P3 , determine the reasonable initial support force F of the leading unit support P .

[0094] Step 5.1: Indoor test to obtain the saturated uniaxial compressive strength of rock R C is 27.3MPa, and the integrity coefficient of the rock mass on the direct roof of the tunnel is K V Take 0.76 as the correction factor K for underground engineering groundwater impact 1 Take 0.2, the correction coefficient K of the main structural surface of underground engineering 2 Take 0, the initial stress state affects the correction factor K 3 Take 1.0, then:

[0095] ;

[0096] .

[0097] The uniaxial compressive strength of rock (coal) obtained from indoor tests is σ c is 30.8MPa, the coal body fracture development coefficient GSI is 85, the mining impact index d is 0.6, and the coal body constant m i Take 6, then:

[0098] ;

[0099] ;

[0100] .

[0101] Step 5.2: Determine the weight coefficient. [BQ] = 251.9, σ = 1.379, according to Table 1, we know that a1 = 0.85, b1 = 0.15.

[0102] Step 5.3: Calculate the initial support force F of the weighted advance bracket P ’ .

[0103] .

[0104] Step 5.4: Compare the weighted leading unit support initial support force F P ’ and the initial support force F of the leading unit support P3 , taking full account of the safety of coal mines, the maximum value is taken as the reasonable initial support force F of the leading unit support P .

[0105] By comparison, we can see that: F P =F P ’ ; That is, the initial support force of the leading unit support is 921.71kN.

[0106] Of course, the above description is not a limitation of the present invention, and the present invention is not limited to the above examples. Changes, modifications, additions or substitutions made by technicians in this technical field within the essential scope of the present invention should also fall within the protection scope of the present invention.

Claims

1. A method for obtaining the initial support force of the advance unit support of a mining tunnel, characterized in that: The following steps are involved: Step 1: Obtain the parameters of the mining tunnel and the advance unit support; The parameters of the mining tunnel and the advance unit support include the elastic-plastic boundary length of the tunnel side, the uniformly distributed load above the rock beam, the length of the rock beam, the bulk density of the overlying rock strata, the distance between the tunnel advance unit support and the coal wall, the ultimate yield strength of the anchor bolt, the tensile strength of the roof rock beam, and the bending section modulus of the rock beam; The method for obtaining the length of the elastic-plastic boundary of the side of the tunnel is as follows: at the leading section of the working face of 150m, a double-end water plug is placed in the middle of the two sides of the tunnel by horizontal drilling. As mining proceeds, at a distance of 120m from the double-end water plug, a measurement is first made at an equal distance of 5m to the borehole, and the water injection loss at different hole depth sections is compared to preliminarily observe the crushing of the coal bodies on the two sides. When the ratio of the front and rear loss is not less than 1.5, the exploration distance is reduced within this section, and the exhaustive method is used to determine that the difference between the current and rear values ​​is within 1m. When the ratio of the front and rear loss is still not less than 1.5, it is considered that the junction of the front section and the rear section is the elastic-plastic boundary of the two sides of the tunnel, that is: （1）； （2）； Where n is the number of injection sections, Q is the amount of injection loss; x i is the distance of each exploration hole of the double-ended water plug; x0 is the elastic-plastic boundary length of the tunnel side; Step 2: Using the parameters of the mining tunnel and the advance unit support, determine the initial support force F of the advance unit support P1 ; The initial support force F of the leading unit support P1 The calculation formula is: （3）； In the formula, is the uniformly distributed load above the rock beam; is the length of the rock beam; is the tensile strength of the roof rock beam; is the bending section modulus of the rock beam; b is the distance between the tunnel advance unit support and the coal wall; Step 3: Determine the steady-state initial support force F of the leading unit support P2 ; The specific steps include: Step 3.1: Obtain the horizontal force transmitted by the two sides of the roadway to the leading unit support , and its calculation formula is as follows: （4）； in, ; A is the pressure coefficient, , μ is Poisson's ratio; S is the size of the base of the advance unit support; K is the stress concentration factor; H is the depth of the coal seam; γ is the bulk density of the overlying rock formation; is the weighted internal friction angle between the direct top and the direct bottom of the roadway; C is the weighted cohesion between the direct top and the direct bottom of the roadway; m is the average thickness of the coal seam; x 实体煤 is the elastic-plastic boundary length of the tunnel side close to the solid coal; x 运输 is the elastic-plastic boundary length of the tunnel side near the belt conveyor; Step 3.2: According to the principle of static balance, the advance unit support can only ensure its own position stability when the horizontal friction force on the advance unit support and the horizontal force generated by the deformation of the roadway coal wall are balanced. That is, the initial support force provided by the advance unit support at this time is the steady-state initial support force. The steady-state initial support force F of the advance unit support is determined according to the following formula: P2 : （5）； In the formula, is the friction force between the top beam of the advance unit support and the top plate of the tunnel; is the friction coefficient of the contact surface between the leading unit bracket and the top plate; Step 4: Determine the initial support force F of the leading unit support P3 ; Determine the initial support force F of the leading unit support according to the following formula P3 ; （6）； In the formula, F P3 is the pressure applied by the advance unit support to the anchor rod, i.e., the initial support force of the advance unit support; S 杆体 is the nominal cross-sectional area of ​​the anchor rod body; σ s is the ultimate yield strength of the anchor rod; Step 5: Carry the initial support force F through the advance unit support P1 , Steady-state initial support force F of the leading unit support P2 and the initial support force F of the leading unit support P3 , determine the reasonable initial support force F of the leading unit support P ; The specific steps include: Step 5.1: Determine the roof equivalent parameter [BQ] and the coal body equivalent parameter σ; The roof equivalent parameter [BQ] is a comprehensive evaluation of the rock mass quality of underground engineering projects, and its expression is as follows: （7）； （8）； Where, BQ is the basic rock mass quality evaluation index; R C is the saturated uniaxial compressive strength of rock; K V is the rock mass integrity index; K1 is the correction coefficient for the influence of groundwater on underground engineering; K2 is the correction coefficient for the influence of the main structural surface of underground engineering; K3 is the correction coefficient for the influence of initial stress state; The coal mass equivalent parameter σ is a comprehensive evaluation of the mechanical properties of the rock mass, and its expression is as follows: （9）； in, , , σ c is the uniaxial compressive strength of coal; GSI is the coefficient of coal fracture development; d is the mining impact index; is a media constant; is the rock hardness coefficient, is the rock crushing coefficient; Step 5.2: Determine the weight coefficient; Determine F based on the roof equivalent parameter [BQ] and the coal body equivalent parameter σ P1 and F P2 The corresponding weight, represented by a1, is F P1 The weight of b1 represents F P2 The weight of When [BQ]>550, σ>10, a1=1, b1=0; When 451<[BQ]≤550, 5<σ≤10, a1=0.95, b1=0.05; When 351<[BQ]≤450, 3<σ≤5, a1=0.9, b1=0.1; When 251<[BQ]≤350, 1<σ≤3, a1=0.85, b1=0.15; When [BQ]≤250, σ≤1, a1=0.8, b1=0.2; Step 5.3: Calculate the initial support force F of the weighted advance bracket P ’ ; The weight factors a1 and b1 form a matrix A; ； The determined leading unit support is subjected to the initial support force F P1 and the steady initial support force F of the leading unit support P2 Composition matrix B; ； Calculate the initial support force F of the weighted advance unit support P ’ ; （10）； Step 5.4: Compare the weighted leading unit support initial support force F P ’ and the initial support force F of the leading unit support P3 , taking full account of the safety of coal mines, the maximum value is taken as the reasonable initial support force F of the leading unit support P ; （11）。

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