Coal seam group mining overlying goaf water inrush risk assessment method

Abstract

The application discloses a coal seam group mining overlying goaf water inrush risk evaluation method and belongs to the technical field of coal mining. The concept of interlayer rock stratum barrier zone is proposed, and the water inrush risk of overlying goaf water is evaluated based on the water blocking performance of the barrier zone. The mechanical structure models of the mining roof and floor are respectively established, the interlayer rock stratum crack distribution characteristics after the coal seam group mining are obtained through a layer-by-layer accumulation and point-by-point judgment calculation method, and the position of the barrier zone is determined. Then, the interlayer rock stratum cracks are introduced into seepage analysis software, and the permeability variation law of the barrier zone is simulated and obtained. Based on the analytic hierarchy process, a multi-factor barrier zone water blocking performance grading evaluation system with the permeability as the core is constructed, an evaluation index of the barrier zone water blocking coefficient is proposed, and the overlying goaf water inrush risk is evaluated. The steps are simple, and the application can provide theoretical guidance for the overlying goaf water prevention and control and water conservation mining of a mine.

Description

Technical Field

[0001] This invention relates to a method for assessing the risk of water accumulation and inrush in the overlying goaf during coal seam mining, belonging to the field of coal mining technology. Background Technology

[0002] Coal, as a sedimentary rock, is mostly found in nature in the form of coal seam groups. During the mining of coal seam groups (mining the upper coal seam first, then the lower coal seam), the interlayer strata are disturbed twice during mining, resulting in deformation, damage, and destruction characteristics that differ from traditional single-seam coal mining. Simultaneously, the goaf of the upper coal seam easily accumulates large amounts of mine water, threatening the safe mining of the lower coal seam.

[0003] The key to preventing water accumulation in the goaf of upper coal seams lies in the water-blocking stability of the inter-layer strata. Previous scholars have conducted extensive research on the water-blocking performance of mined strata, proposing theories such as the aquitard theory, the key aquitard theory, the "barrier-resistance-foundation" theory, and the inter-layer aquitard control layer theory. Currently, intact strata are mostly considered as aquitards, but the water-blocking performance of non-penetrating fractured strata has not received sufficient attention. Especially under coal seam group mining conditions, after the disturbance caused by the superimposed mining of upper and lower coal seams, many inter-layer strata are in a state of either penetration or non-penetration. Therefore, research on the water-blocking performance of non-penetrating fractured strata can provide theoretical support for preventing water accumulation in the goaf of upper coal seams within coal seam groups.

[0004] On the other hand, current technologies for controlling the water-blocking performance of mining-induced rock strata are mostly aimed at preventing fractures, assuming that fractures result in the loss of water-blocking capacity and the absence of fractures indicates effectiveness—a binary 0 / 1 approach. There is a lack of a quantitative "bridge" between the water-blocking performance of rock strata and control technologies, namely, a water-blocking performance evaluation index. Therefore, it is necessary to establish an evaluation system suitable for the water-blocking performance of inter-strata rock strata, and to evaluate the control effect by analyzing the influence of control technologies on water-blocking indices. Summary of the Invention

[0005] Technical Objective: To address the shortcomings of existing technologies, this paper provides a simple and convenient method for evaluating the risk of water accumulation and inrush in the overlying goaf during coal seam mining. The method involves layer-by-layer accumulation and point-by-point judgment of the fracture rate in the mining-induced roof and floor, rapidly determining the location of the water-blocking zone. The interlayer fractures are imported into Comsol seepage analysis software to simulate the permeability variation of the barrier zone. A multi-factor evaluation system for the water-blocking performance of the barrier zone, with permeability as the core, is proposed. Based on the water-blocking coefficient, the risk of water accumulation and inrush in the overlying goaf is graded and evaluated.

[0006] Technical Solution: The present invention provides a method for evaluating the risk of water accumulation and inrush in the overlying goaf during coal seam mining. First, based on actual mining geological conditions, a mechanical model of the elastic foundation combined rock beam of the mining-induced roof and a transversely isotropic planar mechanical model of the mining-induced floor are established. Through layer-by-layer accumulation and point-by-point judgment calculations, the distribution characteristics of fractures in the interlayer rock strata of the coal seam group are obtained, and the specific location of the interlayer barrier zone is determined. Then, the obtained interlayer rock strata fractures are imported into seepage analysis software to simulate the seepage evolution law of the interlayer rock strata under coal seam mining conditions. The permeability of the interlayer rock strata is divided into zones, and the influence of different factors on the seepage of the barrier zone is analyzed. Using the analytic hierarchy process (AHP), a multi-factor evaluation system for the water-blocking performance of the barrier zone is constructed, and a water-blocking coefficient is proposed to quantitatively evaluate the water-blocking performance of the barrier zone. Finally, a grading standard for the water-blocking coefficient of the barrier zone is established to grade and evaluate the risk of water accumulation and inrush in the overlying goaf.

[0007] The specific steps are as follows:

[0008] Step 1: Obtain strata coal and rock samples by field drilling in the study area, conduct mechanical property tests, obtain basic mechanical parameters of the coal seam and the roof and floor strata, and import the basic mechanical parameters of the coal seam and the roof and floor strata into the geological parameter table in Matlab;

[0009] Step 2: Based on the actual mining geological conditions of the study area, establish a mechanical model of the elastic foundation combined rock beam of the mining-induced roof and a mechanical model of the transversely isotropic plane body of the mining-induced floor. Use Matlab to calculate the bending moment and stress distribution of the interlayer strata under the conditions of coal seam group composite mining, which includes mining the upper coal seam first and then mining the lower coal seam. Then, according to the failure judgment criteria, determine the distribution characteristics of all mining-induced fractures in the interlayer strata after coal seam group composite mining, thereby determining the location of the barrier zone.

[0010] Step 3: Calculate the starting coordinates (x1, y1) of the fracture and its corresponding development length l to obtain the ending coordinates (x1, y1+l) of the fracture. Store the starting coordinates (x1, y1) and ending coordinates (x1, y1+l) in two arrays, locate_x and locate_y, respectively. Then, use the line command in Matlab's built-in language to call the starting and ending coordinates in the arrays to generate a fracture distribution map of the mined strata.

[0011] Step 4: Export the fracture distribution map generated in Matlab as a .scr file, then open it with CAD, build a coal seam mining model including the fracture distribution, and generate a .dxf file;

[0012] Step 5: Import the generated .dxf file into Comsol software to build a geological model. Assign corresponding seepage control equations to the fractured and intact rock strata areas in the geological model, start seepage calculation, and obtain the permeability distribution characteristics of the interlayer rock strata after the coal seam group is compositely mined.

[0013] Step Six: Calculate the water blocking coefficient of the barrier zone based on the change in permeability of the barrier zone after the coal seam group is mined, and determine the risk of water accumulation and inrush in the overlying goaf according to the water blocking coefficient classification standard.

[0014] Furthermore, the barrier band is defined as follows:

[0015] The barrier zone refers to the strata in the middle of the inter-seam after the upper and lower coal seams have been mined, where there are intact rock strata without mining-induced fractures and rock strata with non-penetrating fractures. The non-penetrating fracture zone of the floor (roof) and the bending deformation zone of the floor (roof) together constitute the barrier zone. Whether water accumulated in the goaf of the upper coal seam will flow into the working face of the lower coal seam depends crucially on the barrier zone.

[0016] Furthermore, the point-by-point judgment and calculation method is for assessing the development of vertical fractures in the floor after the upper coal seam is mined, and includes the following steps:

[0017] a1. Design the first bottom plate of the upper coal layer as #1, and then number each layer downwards as #2, #3, #i... until the lower coal layer. Establish an xy coordinate system in the interlayer rock between the two coal layers. The origin of the coordinate system is designed in the middle position of the #1 rock layer. The x-axis is positive to the right and the y-axis is positive upward.

[0018] a2. Based on the theory of elastic semi-infinite planar bodies and considering the transverse isotropic characteristics of the bottom rock strata, a mechanical model of the mining bottom plate transverse isotropic semi-infinite planar body is established. The expression for calculating the stress components of the bottom plate is derived, and the stress components (σ) at different coordinate points (x, y) of each rock stratum of the bottom plate are calculated. x σ y ), and convert them to obtain the principal stress components (σ1, σ3);

[0019] a3. Assume we calculate the horizontal coordinate x of the #i-th layer. j Vertical fracture development length l i-j The thickness of this layer is h. i The vertical coordinate of the upper surface of this layer is y. j Then, the coordinates along the thickness direction of this layer can be expressed as (x j y j ), (x j y j -1), (x j y j -2)……(x j yj -s)……(x j y j -h i Substitute the principal stress components (σ1, σ3) at each coordinate point into the maximum tensile stress f. t and Mohr's coulomb shear f s The destruction criteria are as follows: if any one of the destruction criteria is met, the point has been destroyed; if none of the criteria are met, the point has not been destroyed. The horizontal coordinate x of the #i-th layer is determined point by point. j The damage status at each coordinate point along the thickness direction, if coordinate point (x j y j ) to coordinate point (x j y j If both -s) are destroyed, then y is considered to be destroyed. j to y j A fracture was formed at point -s, thus obtaining the x-th rock layer at the horizontal coordinate x. j The length of the vertical fracture development at that location;

[0020] a4. By iterating through step a3, the length of the vertical fracture at any point in the bottom stratum of layer #i can be obtained. Then, by iterating through i = 1, 2, 3..., the distribution of vertical fractures in the entire interlayer stratum after the upper coal seam is mined can be obtained.

[0021] Furthermore, the maximum tensile stress f in step a3 t and Mohr's coulomb shear f s The formula for calculating the failure criterion is:

[0022]

[0023] When f(x,y)≥1, the rock strata in this region are damaged, where, when f t Tensile failure occurs when f ≥ 1. s When the value is ≥1, it indicates shear failure.

[0024] In the formula: c is the cohesion of the rock strata; The internal friction angle of the rock strata.

[0025] Furthermore, the layer-by-layer accumulation calculation method is designed for the development of vertical fractures in the roof after the lower coal seam is mined, and includes the following steps:

[0026] b1. Based on the theory of elastic foundations and combined with actual geological conditions, establish an elastic foundation rock-beam mechanical model and solve for the deflection w of each rock layer in the top plate. i ;

[0027] b2. Assume the perturbation of layer #i+1 is greater than or equal to that of layer #i (w i+1 ≥w iIf the deflection of layer #i and layer #i+1 is less than that of layer #i (w), then layers #i and #i+1 will undergo synchronous bending and subsidence. Similarly, if the deflection of layer #i+1 is less than that of layer #i (w) i+1 <w i At this point, delamination occurs in layers #i and #i+1. Thus, the distribution of delamination fractures can be determined by comparing the deflection deformation between adjacent rock layers. Assuming that there are delamination fractures between layers #m and #m-1, and between #m+2 and #m+3, but no delamination develops between layers #m and #m+2, then layers #m to #m+2 can be regarded as a composite rock beam structure, where m = 2, 3, 4...i-3.

[0028] b3. According to the relevant theory of composite rock beams, the bending moment of the composite beam on any longitudinal section is the sum of the bending moments of each rock layer, and the bending moment M of the composite rock beam can be obtained. Then, the multiple composite beams are decomposed into sub-beams, and the relationship between each sub-beam is confirmed according to the elastic modulus, so as to obtain the distribution of normal stress σ and shear stress τ of each sub-beam.

[0029] b4. Assume we calculate the x-coordinate of the #m-th layer. n The length of the vertical fracture development at the location gradually decreases from the upper surface of the rock beam towards the neutral layer, according to the distribution law of the normal stress in the rock beam; according to the criterion for failure due to maximum tensile stress, when the normal stress σ is greater than the tensile strength [σ t The length of the section is the initial fracture development length l1. Then, the new thickness h is obtained by subtracting the fracture development length from the original rock layer thickness. m -l1; then, through the new rock layer thickness h m -l1 calculates the new distribution of normal stress σ′ in the rock beam. Similarly, by comparing the normal stress and tensile strength, the new fracture development length l2 is obtained. At this time, the remaining thickness of the rock layer becomes h. m -l1-l2; then with the new rock layer thickness h m -l1-l2 calculates the new normal stress distribution σ″, and this calculation continues layer by layer until the final normal stress is less than the tensile strength [σ″]. t Thus, the x-coordinate of the #m layer rock beam is... n The total length of the final vertical fracture development at each location is Σl i If the remaining thickness h of the #m layer rock beam m -Σl i ≤0.1h m At this point, it is assumed that the top slab rock beam of the #m layer is located at the x-coordinate. n At the point contact state, the Mohr-Coulomb shear failure criterion is used to determine whether the crack will break. When τ max When the value is greater than or equal to [τ], it is determined that a fracture has occurred in the rock strata;

[0030] b5. The calculation method in step b4 can be used to obtain the vertical fracture development length of the roof rock beam of layer #m at any horizontal coordinate. Then, by iteratively calculating m = 1, 2, 3..., the fracture distribution of the interlayer rock strata after the lower coal seam is mined can be obtained.

[0031] Furthermore, the formulas for calculating the normal stress σ and shear stress τ of the top strata in step b3 are as follows:

[0032]

[0033] In the formula: M is the equivalent bending moment of the composite beam structure; Q is the equivalent shear force of the composite beam structure; M i σ is the bending moment of the #i-th rock layer; h is the total thickness of the composite rock beam; σ is the normal stress on the longitudinal section of the #m-th rock layer.

[0034] Furthermore, the method for importing inter-layer rock fracture distribution information into seepage analysis software includes the following steps:

[0035] c1. Obtain strata coal and rock samples by field drilling in the study area, conduct mechanical property tests, obtain basic mechanical parameters of coal seam and roof and floor strata, and import the basic mechanical parameters of coal seam and roof and floor strata into the geological parameter table in Matlab.

[0036] c2. Based on the actual mining geological conditions of the study area, a mechanical model of the elastic foundation of the mining-induced roof and a mechanical model of the transversely isotropic plane body of the mining-induced floor were established. The bending moment and stress distribution of the interlayer strata under the condition of coal seam group composite mining were calculated using Matlab. The coal seam group composite mining includes mining the upper coal seam first and then mining the lower coal seam. Then, according to the failure judgment criteria, the development length of all mining-induced fractures in the interlayer strata after coal seam group composite mining was determined, and the coordinate parameters of the fractures were determined at the same time.

[0037] c3. Let the length of the fracture at coordinate (x1, y1) be l. Then the starting and ending coordinates of this vertical fracture are (x1, y1) and (x1, y1+l), respectively. Store the starting and ending coordinates in two arrays, locate_x and locate_y, respectively. Then use the line command in the built-in language of Matlab to call the starting and ending coordinates in the arrays to generate a fracture distribution map of the mined rock strata.

[0038] c4. Export the fracture distribution map generated in Matlab as a .scr file, then open the .scr file with CAD, build a coal seam mining model including the fracture distribution, and generate a .dxf file;

[0039] c5. Import the .dxf file into Comsol software to build a geological model. Assign corresponding seepage control equations to the fractured and intact rock strata regions in the geological model, start seepage calculation, and simulate the permeability variation law of the interlayer barrier zone rock group.

[0040] Furthermore, in the multi-factor evaluation system for the water-blocking coefficient of the barrier strip, the values ​​of each factor are determined as follows:

[0041] ① Lithology C1:

[0042] ②Thickness C2:

[0043] Where: h ij —Original data on the thickness of the barrier zone in the mining area; min(h) ij — The minimum value in the original barrier band thickness data; max(h) ij — The maximum value in the original barrier band thickness data; C2 is the data after normalization.

[0044] ③ Rock strata assemblage C3:

[0045] ④ Penetration rate C4:

[0046] In the formula: k is the equivalent permeability of the barrier zone.

[0047] ⑤Hydraulic properties C5: The main hydraulic property of rocks that significantly affects their water-blocking performance is expansibility T. p and disintegration T b C5 = T p -T b |,

[0048] in:

[0049] ⑥ C6 mining height:

[0050] Where: M ij This provides the raw data for mining height in the mining area; min(M) ij ) represents the minimum value in the sampling height data; max(M) ij () represents the maximum value in the height data;

[0051] ⑦ Working face dip length C7:

[0052] In the formula: L ij This provides the raw data for mining height in the mining area; min(L) ij ) represents the minimum value in the sampling height data; max(L) ij() represents the maximum value in the height data.

[0053] Furthermore, the method for calculating the water-blocking coefficient ZI of the barrier strip is as follows:

[0054] In the formula: ZI is the water resistance coefficient; W i The weights of each factor affecting the water resistance coefficient are: f i (x,y) represents the normalized values ​​of various influencing factors of the water resistance coefficient.

[0055] Furthermore, the classification criteria for the water-blocking coefficient are as follows: ZI < 0.36 indicates a high risk of water inrush, 0.36 ≤ ZI < 0.54 indicates a medium risk of water inrush, 0.54 ≤ ZI < 0.7 indicates a low risk of water inrush, and ZI ≥ 0.7 indicates no risk of water inrush.

[0056] Beneficial effects:

[0057] 1) This method proposes the concept of a barrier zone, which regards both intact rock strata and rock strata containing non-penetrating fractures as water-blocking rock groups, and uses their water-blocking performance as an evaluation index for water accumulation and inrush in the overlying goaf, providing a new approach for the prevention and control of water inrush in the overlying water body.

[0058] 2) This method proposes a method for calculating the layer-by-layer accumulation of mining-induced roof fractures and a method for calculating the point-by-point judgment of mining-induced floor fractures. This method overcomes the shortcomings of traditional calculation methods that only consider whether the rock strata have broken, and truly calculates the actual development length of mining-induced fractures.

[0059] 3) This method proposes a way to import fractures into seepage analysis software, which realizes the combination of actual mining-induced rock strata damage with seepage calculation, and overcomes the situation that traditional mining-induced rock strata seepage calculation does not consider fracture seepage.

[0060] 4) This method proposes an evaluation index for the water-blocking performance of barrier strips, which comprehensively considers a variety of influencing factors and realizes a quantitative evaluation of the water-blocking performance of barrier strips. Attached Figure Description

[0061] Figure 1 This is a flowchart illustrating the method for assessing the risk of water accumulation and inrush in the overlying goaf during coal seam mining according to the present invention.

[0062] Figure 2 This is a schematic diagram showing the location of the barrier strip in this invention.

[0063] Figure 3 This is a schematic diagram of the transversely isotropic semi-infinite planar body model of the moving base plate of the present invention.

[0064] Figure 4 This is a schematic diagram illustrating the calculation of the development length of the crack in the mining base plate according to the present invention.

[0065] Figure 5 This is a schematic diagram of the elastic foundation beam model of the mining roof of the present invention.

[0066] Figure 6 This is a schematic diagram of the mining-driven roof composite rock beam of the present invention.

[0067] Figure 7 This is a schematic diagram illustrating the calculation of the development length of the mining-induced roof cracks according to the present invention.

[0068] Figure 8 This is a schematic diagram of the calculation method of layer-by-layer accumulation and point-by-point judgment in this invention.

[0069] Figure 9 This is a schematic diagram of the distribution of interlayer rock fractures exported from CAD in this invention.

[0070] Figure 10 This is a schematic diagram of seepage distribution in the geological model constructed by Comsol in this invention.

[0071] Figure 11 This is a schematic diagram of the hierarchical analysis method for evaluating the water-blocking performance of the barrier strip according to the present invention.

[0072] Figure 12 This is a schematic diagram illustrating the risk assessment of water accumulation and inrush in the overlying goaf area according to the present invention.

[0073] In the diagram: 1-Upper coal seam, 2-Upper coal seam goaf, 3-Inter-layer strata, 4-Lower coal seam, 5-Lower coal seam goaf, 6-Non-penetrating fracture, 7-Penetrating fracture, 8-Penetrating fracture zone of the floor, 9-Non-penetrating fracture zone of the floor, 10-Roof (floor) bending deformation zone, 11-Non-penetrating fracture zone of the roof, 12-Penetrating fracture zone of the roof, 13-Roof collapse zone, 14-Barrier zone. Detailed Implementation

[0074] The embodiments of the present invention will be further described below with reference to the accompanying drawings:

[0075] like Figure 1 , Figure 2 As shown, this method presents a method for calculating the fracture rate of the mining-induced roof and floor layers by accumulating layer by layer and judging point by point, which can quickly determine the location of the barrier zone with water-blocking function. The interlayer fractures are imported into the Comsol seepage analysis software to simulate the permeability variation law of the barrier zone. A multi-factor evaluation system for the water-blocking performance of the barrier zone with permeability as the core is proposed. The risk of water accumulation and inrush in the overlying goaf is graded and evaluated based on the water-blocking coefficient.

[0076] Based on the geological conditions of a certain coal mine, the upper coal seam 1 has been mined out, forming a large-scale goaf 2 with a large accumulation of mine water. Therefore, after the lower coal seam 4 is mined, a lower coal seam goaf 5 is formed. During the mining process and the collapse of the goaf, the following fracture zones are formed: floor through fracture zone 8, floor non-through fracture zone 9, roof (floor) bending deformation zone 10, roof non-through fracture zone 11, roof through fracture zone 12, roof collapse zone 13, and barrier zone 14, including two types of fractures: non-through fracture 6 and through fracture 7. During the mining process, the mine is threatened by water accumulation in the overlying goaf 2, so it is necessary to analyze the risk of water inrush during the mining process.

[0077] like Figure 1 As shown, the specific implementation steps of the method for assessing the risk of water accumulation and inrush in the overlying goaf of coal seam mining according to the present invention are as follows:

[0078] Step 1: Drill the interlayer rock strata 3 to obtain coal and rock samples, perform mechanical property tests, obtain the basic mechanical parameters of the coal seam and the roof and floor rock strata, and import the basic mechanical parameters of the coal seam and the roof and floor rock strata into the geological parameter table in Matlab;

[0079] Step 2: Utilize the layer-by-layer accumulation and point-by-point judgment algorithm developed in Matlab ( Figure 8 The bending moment and stress distribution of the interlayer strata under the combined mining conditions of coal seam group were calculated respectively. The combined mining of coal seam group includes mining the upper coal seam first and then mining the lower coal seam 4. Then, according to the failure judgment criteria, the distribution characteristics of all mining-induced fractures in the interlayer strata after the combined mining of coal seam group were determined, thereby determining the location of the barrier zone 14.

[0080] like Figure 8 As shown, the point-by-point judgment calculation method is for the development of vertical fractures in the floor after the upper coal seam is mined, and includes the following steps:

[0081] a1. Design the first bottom plate of the upper coal layer as #1, and then number each layer downwards as #2, #3, #i... until the lower coal layer. Establish an xy coordinate system in the interlayer rock between the two coal layers. The origin of the coordinate system is designed in the middle position of the #1 rock layer. The x-axis is positive to the right and the y-axis is positive upward.

[0082] a2. Based on the theory of elastic semi-infinite planar bodies and considering the transverse isotropic characteristics of the bottom rock strata, establish a mechanical model of a transversely isotropic semi-infinite planar body for mining operations, such as... Figure 3 As shown, the stress component calculation expression of the floor plate is derived using the transverse isotropic semi-infinite plane mechanical model of the mining floor plate. The stress components (σ) at different coordinate points (x, y) of each rock layer of the floor plate are calculated. x σ y ), and convert them to obtain the principal stress components (σ1, σ3), such as Figure 4As shown;

[0083] a3. Let the horizontal coordinate of the #i-th layer be x. j Vertical fracture development length l i-j The thickness of this layer is h. i The vertical coordinate of the upper surface of this layer is y. j Then, the coordinates along the thickness direction of this layer can be expressed as (x j y j ), (x j y j -1), (x j y j -2)……(x j y j -s)……(x j y j -h i Substitute the principal stress components (σ1, σ3) at each coordinate point into the maximum tensile stress f. t and Mohr's coulomb shear f s The destruction criteria are as follows: if any one of the destruction criteria is met, the point is considered destroyed; if none of the criteria are met, the point is considered not destroyed. The horizontal coordinate x of the #i-th layer is determined point by point. j The damage status at each coordinate point along the thickness direction, if coordinate point (x j y j ) to coordinate point (x j y j If both -s) are destroyed, then y is considered to be destroyed. j to y j A fracture was formed at point -s, thus obtaining the x-th rock layer at the horizontal coordinate x. j The length of the vertical fracture development at that location;

[0084] a4. By iterating through step a3, the length of the vertical fracture at any point in the bottom stratum of layer #i can be obtained. Then, by iterating through i = 1, 2, 3..., the distribution of vertical fractures in the entire interlayer stratum after the upper coal seam is mined can be obtained.

[0085] Maximum tensile stress f t and Mohr's coulomb shear f s The formula for calculating the failure criterion is:

[0086]

[0087] When f(x,y)≥1, the rock strata in this region are damaged. Specifically, when f... t Tensile failure occurs when f ≥ 1. s When the value is ≥1, it indicates shear failure.

[0088] like Figure 8 As shown, the layer-by-layer accumulation calculation method is based on the development of vertical fractures in the roof after the mining of the lower coal seam 4, and includes the following steps:

[0089] b1. Based on the theory of elastic foundations and combined with actual geological conditions, establish a mechanical model of elastic foundation rock beams, such as... Figure 5 As shown, solve for the deflection w of each rock layer in the top plate. i ;

[0090] b2. Assume the perturbation of layer #i+1 is greater than or equal to that of layer #i (w i+1 ≥w i If the deflection of layer #i and layer #i+1 is less than that of layer #i (w), then layers #i and #i+1 will undergo synchronous bending and subsidence. Similarly, if the deflection of layer #i+1 is less than that of layer #i (w) i+1 <w i At this point, delamination occurs in layers #i and #i+1. Thus, the distribution of delamination fractures can be determined by comparing the deflection deformation between adjacent rock layers. Assuming that delamination fractures exist between layers #m and #m-1, and between #m+2 and #m+3, but not between layers #m and #m+2, then layers #m to #m+2 can be considered as a composite rock beam structure. Figure 6 As shown, m = 2, 3, 4...i-3;

[0091] b3. According to the relevant theory of composite rock beams, the bending moment of the composite beam at any longitudinal section is the sum of the bending moments of each rock layer, and the bending moment M of the composite rock beam can be obtained. Then, the multiple composite beams are decomposed into sub-beams, and the relationship between each sub-beam is determined according to the elastic modulus, so as to obtain the distribution of normal stress σ and shear stress τ of each sub-beam, such as... Figure 7 As shown;

[0092] The formulas for calculating the normal stress σ and shear stress τ of the top stratum are as follows:

[0093]

[0094] In the formula: M is the equivalent bending moment of the composite beam structure; Q is the equivalent shear force of the composite beam structure; M i σ is the bending moment of the #i-th rock layer; h is the total thickness of the composite rock beam; σ is the normal stress on the longitudinal section of the #m-th rock layer.

[0095] b4. Assume we calculate the x-coordinate of the #m-th layer. n The length of the vertical fracture development at the location gradually decreases from the upper surface of the rock beam towards the neutral layer, according to the distribution law of the normal stress in the rock beam; according to the criterion for failure due to maximum tensile stress, when the normal stress σ is greater than the tensile strength [σ t The length of the section is the initial fracture development length l1. Then, the new thickness h is obtained by subtracting the fracture development length from the original rock layer thickness. m-l1; then, through the new rock layer thickness h m -l1 calculates the new distribution of normal stress σ′ in the rock beam. Similarly, by comparing the normal stress and tensile strength, the new fracture development length l2 is obtained. At this time, the remaining thickness of the rock layer becomes h. m -l1-l2; then with the new rock layer thickness h m -l1-l2 calculates the new normal stress distribution σ″, and this calculation continues layer by layer until the final normal stress is less than the tensile strength [σ″]. t Thus, the x-coordinate of the m-layer rock beam is... n The total length of the final vertical fracture development at each location is Σl i If the remaining thickness h of the m-layer rock beam m -Σl i ≤0.1h m At this point, it is assumed that the top slab rock beam of the #m layer is located at the x-coordinate. n At the point contact state, the Mohr-Coulomb shear failure criterion is used to determine whether the crack will break. When τ max When the value is greater than or equal to [τ], it is determined that a fracture has occurred in the rock strata;

[0096] b5. The calculation method in step b4 can be used to obtain the vertical fracture development length of the roof rock beam of layer #m at any horizontal coordinate. Then, by iteratively calculating m = 1, 2, 3..., the fracture distribution of the interlayer rock strata after the mining of the lower coal seam 4 can be obtained.

[0097] Step 3: Let the length of the fracture at coordinate (x1, y1) be l. Then the starting and ending coordinates of this vertical fracture are (x1, y1) and (x1, y1+l), respectively. Store the starting and ending coordinates in two arrays, locate_x and locate_y, respectively. Then use the line command in the Matlab built-in language to call the starting and ending coordinates in the arrays to generate a fracture distribution map of the mined rock strata.

[0098] Step 4: Export the fracture distribution map generated in Matlab as a .scr file, then open the .scr file with CAD, build a coal seam mining model including the fracture distribution, and export it as a .dxf file;

[0099] Step 5: Import the .dxf file into Comsol software to build a geological model. Assign corresponding seepage control equations to the fractured and intact rock strata areas in the geological model, start seepage calculation, and obtain the permeability distribution characteristics of the interlayer rock strata after the coal seam group is compositely mined.

[0100] like Figure 9 , Figure 10 As shown, the specific method for importing interlayer rock fracture distribution information into seepage analysis software is as follows:

[0101] c1. Obtain strata coal and rock samples by field drilling in the study area, conduct mechanical property tests, obtain basic mechanical parameters of coal seam and roof and floor strata, and import the basic mechanical parameters of coal seam and roof and floor strata into the geological parameter table in Matlab.

[0102] c2. Based on the actual mining geological conditions of the study area, establish a transverse isotropic planar mechanical model of the mining-induced floor and a combined rock-beam mechanical model of the mining-induced roof elastic foundation, such as... Figure 3 and Figure 5 As shown, the bending moment and stress distribution of inter-layer strata under the condition of coal seam group composite mining were calculated using Matlab. The coal seam group composite mining includes mining the upper coal seam first and then mining the lower coal seam 4. Then, according to the failure judgment criteria, the development length of all mining-induced fractures in the inter-layer strata after coal seam group composite mining was determined.

[0103] c3. Let the length of the fracture development at coordinate (x1, y1) be l, such as Figure 7 As shown, the starting and ending coordinates of the vertical fracture are (x1, y1) and (x1, y1+l), respectively. The starting and ending coordinates are stored in two arrays, locate_x and locate_y, respectively. Then, the line command in the Matlab built-in language is used to call the starting and ending coordinates in the arrays to generate a fracture distribution map of the mined rock layer.

[0104] c4. Export the fracture distribution map generated in Matlab as a .scr file, then open the .scr file with CAD, build a coal seam mining model including the fracture distribution, and generate a .dxf file;

[0105] c5. Import the .dxf file into Comsol software to build a geological model. Assign corresponding seepage control equations to the fractured and intact rock strata regions in the geological model, start seepage calculation, and simulate the permeability variation law of the interlayer barrier zone rock group.

[0106] Step Six: Establish a water-blocking performance rating system for the barrier strip based on the analytic hierarchy process (AHP), such as... Figure 11 As shown, the water-blocking coefficient of the barrier zone is calculated, and the risk of water accumulation and inrush in the overlying goaf is determined according to the water-blocking coefficient grading standard. Figure 12 As shown.

[0107] A multi-factor evaluation system for the water-blocking coefficient of the barrier zone is established. The multi-factor evaluation system for the water-blocking coefficient of the barrier zone has a tree structure. The first layer is the water-blocking coefficient A, which reflects the water-blocking performance of the barrier zone. The second layer is the criterion layer, which includes engineering geological conditions B1, hydrogeological conditions B2, and mining conditions B3. The third layer is the sub-criterion layer, which includes lithology C1, thickness C2, and rock strata combination C3 associated with engineering geological conditions B1, permeability C4 and hydrogeological properties C5 associated with hydrogeological conditions B2, and mining height C6 and working face dip length C7 associated with mining conditions B3.

[0108] In the aforementioned water-blocking coefficient evaluation system for the barrier zone, the weights of each influencing factor are determined by inviting experts in relevant fields within the industry to score the influence weights of each factor. Then, the collected data are summarized to form a judgment matrix, and a consistency test is performed. The influencing factors include: lithology C1, thickness C2, rock strata combination C3, permeability C4, hydrogeological properties C5, mining height C6, and working face dip length C7. The corresponding weights of each influencing factor are: W1 = 0.0740, W2 = 0.1762, W3 = 0.0466, W4 = 0.1349, W5 = 0.4047, W6 = 0.1089, and W7 = 0.0545.

[0109] The multi-factor evaluation system for the water-blocking coefficient of the barrier zone is established, and the values ​​of each factor are determined as follows:

[0110] ① Lithology C1:

[0111] ②Thickness C2:

[0112] Where: h ij —Original data on the thickness of the barrier zone in the mining area; min(h) ij — The minimum value in the original barrier band thickness data; max(h) ij — The maximum value in the original barrier band thickness data; C2 is the data after normalization.

[0113] ③ Rock strata assemblage C3:

[0114] ④ Penetration rate C4:

[0115] In the formula: k is the equivalent permeability of the barrier zone.

[0116] ⑤Hydraulic properties C5: The main hydraulic property of rocks that significantly affects their water-blocking performance is expansibility T. p and disintegration T b C5 = T p -T b |,

[0117] in:

[0118] ⑥ C6 mining height:

[0119] Where: M ij This provides the raw data for mining height in the mining area; min(M) ij ) represents the minimum value in the sampling height data; max(M) ij () represents the maximum value in the height data;

[0120] ⑦ Working face dip length C7:

[0121] In the formula: L ij This provides the raw data for mining height in the mining area; min(L) ij ) represents the minimum value in the sampling height data; max(L) ij () represents the maximum value in the height data.

[0122] The method for calculating the water-blocking coefficient ZI of the barrier zone is as follows:

[0123] In the formula: ZI is the water resistance coefficient; W i The weights of each factor affecting the water resistance coefficient are: f i (x, y) represents the normalized values ​​of various influencing factors of the water-blocking coefficient. The aforementioned grading and evaluation criteria for the water-blocking coefficient of the barrier zone include: ZI < 0.36 indicates high risk of water inrush, 0.36 ≤ ZI < 0.54 indicates medium risk of water inrush, 0.54 ≤ ZI < 0.7 indicates low risk of water inrush, and ZI ≥ 0.7 indicates no risk of water inrush.

Claims

1. A method for evaluating the risk of water inrush from water in goaf overlying coal seams, characterized in that: The complete rock stratum without mining-induced fracture development in the middle of interlayer rock strata and the rock stratum with non-penetrating fracture development are collectively referred to as a barrier zone under the condition of coal seam group mining, and the water-blocking performance of the barrier zone is used as an index for evaluating the water inrush risk of the water accumulated in the overlying goaf; According to the actual mining geological information, a transversely isotropic plane body mechanics model of the mining floor and an elastic foundation combined rock beam mechanics model of the mining roof are respectively established; The point-by-point judgment calculation method is used for the transversely isotropic plane body mechanics model of the mining floor to calculate the fracture development of the interlayer rock strata after the upper coal seam is mined, and the layer-by-layer accumulation calculation method is used for the elastic foundation combined rock beam mechanics model of the mining roof to calculate the fracture development of the interlayer rock strata after the lower coal seam is mined, so as to obtain the fracture distribution characteristics of the interlayer rock strata after the upper and lower coal seams are mined and determine the specific position of the barrier zone; In the point-by-point judgment calculation method for the transversely isotropic plane body mechanics model of the mining floor: a1, the first floor of the upper coal seam is numbered as #1, and then the floor is numbered as #2, #3, #i, and so on downward until the lower coal seam, an x-y coordinate system is established for the interlayer rock strata between the two coal seams, the origin of the coordinate system is designed at the middle position of the #1 rock stratum, the x-axis is positive to the right, and the y-axis is positive upward; a2, based on the elastic semi-infinite plane body theory and considering the transverse isotropic characteristics of the floor strata, a mining floor transverse isotropic semi-infinite plane body mechanical model is established, the floor stress component calculation expression is derived by using the mining floor transverse isotropic semi-infinite plane body mechanical model, the stress components (σ x , σ y ) at different coordinate points (x, y) of each floor stratum are calculated, and the principal stress components (σ1, σ3) are converted; a3, the horizontal coordinate of the i-th layer is x j the length of vertical fracture development at x i-j , the thickness of the layer is h i , the longitudinal coordinate of the upper surface of the layer is y j , the coordinates of the layer along the thickness direction are respectively represented as (x j , y j ), (x j , y j -1), (x j , y j -2)…(x j , y j -s)…(x j , y j -h i ); the principal stress components (σ1, σ3) of each coordinate point are substituted into the maximum tensile stress f t and the Mohr-Coulomb shear f s failure criterion, if any of the failure criteria is met, it is judged that the point has failed, if none of the failure criteria is met, it is judged that the point has not failed; the failure of each coordinate point along the thickness direction at the horizontal coordinate x j of the i-th layer is judged, if the coordinate point (x j , y j ) to the coordinate point (x j , y j -s) all fail, it is considered that a fracture is formed at y j to y j -s, thus obtaining the length of vertical fracture development at the horizontal coordinate x j of the i-th layer The specific position information of the interlayer rock stratum fracture obtained is introduced into a seepage analysis software, and the seepage evolution of the interlayer rock strata under the condition of coal seam group mining is simulated by using the seepage analysis software to obtain the permeability variation law of the barrier zone; An evaluation system of the water-blocking performance of the barrier zone with permeability as the core is constructed by using the analytic hierarchy process, and a water-blocking coefficient is used to quantitatively evaluate the water-blocking performance of the barrier zone; According to the natural breakpoint method, a barrier zone water-blocking coefficient grading standard is established to classify and evaluate the water inrush risk of the water accumulated in the overlying goaf when the lower coal seam is mined.

2. The method according to claim 1, wherein, The point-by-point judgment calculation method is used for the transversely isotropic plane body mechanics model of the mining floor to calculate the fracture development of the interlayer rock strata after the upper coal seam is mined, and the layer-by-layer accumulation calculation method is used for the elastic foundation combined rock beam mechanics model of the mining roof to calculate the fracture development of the interlayer rock strata after the lower coal seam is mined, so as to obtain the fracture distribution characteristics of the interlayer rock strata after the upper and lower coal seams are mined and determine the specific position of the barrier zone; The layer-by-layer accumulation calculation method is used for the vertical fracture development of the roof after the lower coal seam is mined, including the following steps:

3. The method according to claim 2, wherein, b3, according to the related theory of combined rock beam, the bending moment of the combined beam on an arbitrary longitudinal section is the sum of the bending moments of each rock stratum, the bending moment M of the combined rock beam is obtained; then, the multiple combined beams are decomposed into sub-beams, and the relationship between each sub-beam is determined according to the elastic modulus to obtain the normal stress σ and shear stress τ distribution of each sub-beam; b1、According to the elastic foundation theory, an elastic foundation beam mechanical model is established combined with the actual geological conditions to solve the perturbation w of each rock stratum of the roof i ; b2, if the disturbance of the (i+1)th layer is greater than or equal to the ith layer (w i+1 ≥w i ), then the ith and (i+1)th layers are in synchronous bending subsidence; similarly, if the disturbance of the (i+1)th layer is less than the ith layer (w i+1 <w i ), then the ith and (i+1)th layers are in separation, so that the distribution of separation fissures is determined by comparing the disturbance deformations between adjacent rock layers; it is assumed that there are separation fissures between the mth and (m-1)th layers and between (m+2)th and (m+3)th layers, but no separation fissures between the mth and (m+2)th layers, then the mth to (m+2)th layers are regarded as a combined rock beam structure, m=2, 3, 4, …, i-3; b5, the vertical fracture development length of the #m layer roof rock beam at an arbitrary transverse coordinate is obtained by using the calculation method in step b4, and then the fracture distribution of the interlayer rock strata after the lower coal seam is mined is obtained by cyclic calculation m=1, 2, 3… b4. Assume that the horizontal coordinate x of the mth layer is calculated n The length of vertical crack development at the position of x is gradually reduced from the upper surface of the rock beam to the neutral layer according to the distribution of the normal stress of the rock beam. According to the maximum tensile stress failure criterion, the length of the section where the normal stress σ is greater than the tensile strength [σ t ] is the initial crack development length l1. Then, the original rock layer thickness is reduced by the crack development length to obtain a new thickness h m -l1. Then, the new normal stress σ' distribution of the rock beam is calculated by using the new rock layer thickness h m -l1. Similarly, the new crack development length l2 is obtained by comparing the normal stress and the tensile strength. At this time, the remaining thickness of the rock layer becomes h m -l1-l2. The new normal stress distribution σ'' is calculated by using the new rock layer thickness h m -l1-l2. The calculation is performed layer by layer until the normal stress is less than the tensile strength [σ t ]. The total length of the final vertical crack development at the position of the horizontal coordinate x n of the mth layer rock beam is Σl i . If the remaining thickness h m -Σl i of the mth layer rock beam is less than or equal to 0.1h m , it is considered that the mth layer rock beam is in a point contact state at the position of the horizontal coordinate x n . According to the Mohr-Coulomb shear failure criterion, it is determined whether the crack is broken when τ max ≥[τ]. The normal stress σ and shear stress τ calculation formula of the roof rock stratum in step b3 is:

4. The method according to claim 2, characterized in that, The maximum tensile stress f in step a3 t And the Mohr-Coulomb shear f s The failure criterion calculation formula is: When f(x, y) > 1, the rock formation in the region fails, where, when f t > 1 is tensile failure, and when f s > 1 is shear failure. where: c is the cohesion of the rock formation; is the internal friction angle of the rock formation.

5. The coal seam group mining overlying goaf water inrush water hazard evaluation method according to claim 3, characterized in that, The method of introducing the fracture distribution information of the interlayer rock strata into the seepage analysis software is as follows: where: M is the equivalent bending moment of the composite beam structure; Q is the equivalent shear of the composite beam structure; M i is the bending moment of the ith layer of rock; h is the total thickness of the composite rock beam; and σ is the normal stress on the longitudinal cross-section of the mth layer of rock.

6. The coal group mining overlying goaf water inrush danger assessment method of claim 1, characterized in that, ​ c1, field drilling in the study area obtains coal rock samples, mechanical property tests are performed, and basic mechanical parameters of the coal seam and the roof and floor strata are obtained, which are introduced into the geological parameter table in Matlab; c2, according to the actual mining geological conditions of the study area, an elastic foundation combined rock beam mechanical model of the mining roof and a transversely isotropic plane body mechanical model of the mining floor are respectively established, and the bending moment and stress distribution of the interlayer strata under the condition of composite mining of the coal seam group are calculated by using Matlab, wherein the composite mining of the coal seam group includes mining the upper coal seam first and then mining the lower coal seam; then, according to the failure criterion, the development length of all mining-induced fractures in the interlayer strata after composite mining of the coal seam group is determined; c3, assuming that the fracture development length at the coordinate (x1, y1) is l, the starting point coordinate and the end point coordinate of the vertical fracture are (x1, y1) and (x1, y1+l) respectively; the starting point coordinate and the end point coordinate are saved in two arrays locate_x and locate_y respectively; then, the starting point and the end point coordinates in the arrays are called by using the line command in the built-in language of Matlab, and the fracture distribution diagram of the mining strata is generated; c4, the fracture distribution diagram generated in Matlab is exported as a.scr file, then the.scr file is opened by CAD, and a coal seam mining model including the fracture distribution is established, and a.dxf file is generated; c5, the.dxf file is imported into Comsol software to construct a geological model, the corresponding seepage control equation is given to the fractures and the complete rock strata in the geological model respectively, the seepage calculation is started, and the variation law of the permeability of the interlayer barrier zone rock group is simulated.

7. The coal group mining overlying goaf water inrush danger assessment method of claim 1, characterized in that: A multi-factor evaluation system of barrier zone water resistance coefficient is established, which is a tree structure, wherein the first layer is a water resistance coefficient A reflecting the water resistance performance of the barrier zone, the second layer is a criterion layer, including engineering geological conditions B1, hydrogeological conditions B2 and mining conditions B3; the third layer is a sub-criterion layer, including lithology C1, thickness C2 and rock combination C3 associated with the engineering geological conditions B1, permeability C4 and water properties C5 associated with the hydrogeological conditions B2, mining height C6 and working face inclination length C7 associated with the mining conditions B3; In the evaluation system of the barrier zone water resistance coefficient, the influence weights of various factors are determined by inviting experts in the relevant fields to score, then the collected data are summarized to form a judgment matrix, and consistency test is performed; The influence factors include: lithology C1, thickness C2, rock combination C3, permeability C4, water properties C5, mining height C6, working face inclination length C7, and the corresponding weights of each influence factor are: W1=0.0740, W2=0.1762, W3=0.0466, W4=0.1349, W5=0.4047, W6=0.1089, W7=0.0545.

8. The method according to claim 7, characterized in that: The multi-factor evaluation system of the barrier zone water resistance coefficient is determined as follows: ①Lithology C1: (ii) thickness C2: where: h ij — thickness of the mine barrier zone raw data; min(h ij — minimum value in the raw barrier zone thickness data; max(h ij — maximum value in the raw barrier zone thickness data; C2 is the normalized data; iii. Rock combination C3: (4) permeability C4: where: k is the equivalent permeability of the barrier tape, (5) Water physical property C5: The water physical property of rock that has a large influence on water blocking performance is expansibility T p and disintegration T b , C5 = |T p - T b |, wherein: C6: where: M ij is the raw data of the mining area's height; min(M ij ) is the minimum value in the data of the mining area's height; max(M ij ) is the maximum value in the data of the mining area's height; • Length of the working face in the direction of the dip C7: where: L ij is the raw data of the mining area's height; min(L ij ) is the minimum value in the height data; and max(L ij ) is the maximum value in the height data.

9. The method according to claim 8, characterized in that: The calculation method of the barrier zone water resistance coefficient ZI is: In the formula, ZI is the water resistance coefficient; W i is the weight of each influencing factor of the water resistance coefficient; f i (x, y) is the normalized value of each influencing factor of the water resistance coefficient.

10. The method according to claim 9, wherein the method further comprises: determining the water inrush risk of the water in the goaf of the overlying coal seam group. The blocking band water blocking coefficient grading evaluation standard includes that ZI < 0.36 belongs to high water breakthrough risk, 0.36 <= ZI < 0.54 belongs to medium water breakthrough risk, 0.54 <= ZI < 0.7 belongs to low water breakthrough risk, and ZI >= 0.7 belongs to no water breakthrough risk.

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