Method for determining load of surrounding rock of adjacent gob roadway of coal mine and judging impact instability of surrounding rock

Abstract

The invention discloses a coal mine adjacent goaf roadway surrounding rock load determination and impact instability judgment method, which comprises the following steps: establishing an adjacent goaf working face initial pressure and full mining stage working face roof structure model, an adjacent goaf working face initial pressure and full mining stage working face roof three-dimensional model and adjacent goaf roadway three-load superposition schematic; the adjacent goaf roadway surrounding rock load-bearing characteristics and the roof fracture load transfer rule of the adjacent goaf working face at different mining stages are analyzed; a method for calculating lateral static load stress of an adjacent goaf, advanced static load stress of the adjacent goaf and fracture dynamic load of an overlying thick and hard rock stratum borne by surrounding rock of the adjacent goaf roadway in different mining stages of the adjacent goaf working face is provided, and the load bearing state of the surrounding rock of the adjacent goaf roadway is quantitatively represented; according to the method, the adjacent goaf surrounding rock impact instability criterion is provided, the adjacent goaf surrounding rock stability is theoretically analyzed, the adjacent goaf rock burst prevention and control thought and method are analyzed based on the impact instability criterion, and the method has important theoretical and practical significance on adjacent goaf rock burst prevention and control.

Description

Technical Field

[0001] This invention relates to the field of coal mine safety technology, and in particular to a method for determining the surrounding rock load in coal mine adjacent roadways and identifying its impact instability. Background Technology

[0002] Coal, as a crucial pillar of my country's energy supply, faces increasingly prominent rockburst problems during deep mining, becoming a key bottleneck restricting the safe and efficient exploitation of coal resources. In related technologies, a roadway surrounding rock stability assessment system has been constructed through the collaborative operation of geomechanical analysis, stress monitoring, and support design. Specifically, this system covers the entire process from rock strata structure identification and surrounding rock load analysis to the formulation of impact instability criteria, including key aspects such as critical layer fracture law analysis, dynamic and static load superposition calculation, and roadway surrounding rock instability analysis. With increasing mining depth and frequency of mining disturbances, the stress environment of the surrounding rock in adjacent workings becomes increasingly complex. While traditional methods have some applicability in shallow mining, they are insufficient to accurately characterize the multi-stage load superposition effect and impact instability mechanism of the surrounding rock in adjacent workings under conditions of thick, hard roofs in deep mining.

[0003] However, existing methods for calculating roadway surrounding rock loads typically employ a single static load model for stress analysis, failing to fully consider the synergistic effects of lateral static loads from adjacent goaf areas, pre-existing static loads from the working face, and dynamic loads from the fracturing of the overlying thick, hard rock strata. This can lead to biased load assessments or inaccurate instability detection, thus affecting the prevention and control of rockbursts. Especially during the initial pressure and full mining stages of adjacent goaf faces, the roof structure morphology varies significantly, and current technologies lack detailed load modeling and dynamic response analysis for different mining stages, making it difficult to scientifically predict and effectively intervene in roadway rockburst risks. Summary of the Invention

[0004] The present invention aims to at least partially solve one of the technical problems in the related art.

[0005] Therefore, the first objective of this invention is to propose a method for determining the surrounding rock load and identifying its impact instability in coal mine adjacent roadways.

[0006] Another objective of this invention is to provide a device for determining the load on the surrounding rock of a coal mine adjacent roadway and for identifying its impact instability.

[0007] To achieve the above objectives, this invention proposes a method for determining the surrounding rock load and identifying its impact instability in coal mine adjacent working roadways, comprising: S1. Based on the borehole columnar section near the working face, a roof structure model is established for the initial pressure and full mining stages, and the loading characteristics of the surrounding rock in the roadway at different mining stages are analyzed. S2 calculates the lateral static stress of the adjacent goaf, the advanced static stress of the adjacent working face, and the dynamic stress caused by the fracture of the overlying thick and hard rock layer on the surrounding rock of the roadway, and realizes the total stress calculation under the superposition of the three loads. S3, based on the difference between the total stress on the surrounding rock of the roadway and the roadway support stress, combined with the critical stress of coal and rock mass impact instability, calculate the impact instability index of the surrounding rock of the roadway, which is used to determine whether the roadway has experienced impact instability. S4. Based on the calculation results of the rock impact instability index of the roadway surrounding rock, measures for preventing roadway rockburst are proposed, including using high-strength rockburst-resistant anchor bolts or constant resistance energy-absorbing anchor bolts for roadway support, or implementing hydraulic fracturing of the roof area and pre-fracturing blasting to relieve pressure in the advanced area of ​​the working face.

[0008] In addition, the method for determining the surrounding rock load and identifying its impact instability in a coal mine adjacent roadway according to the above embodiments of the present invention may also have the following additional technical features: Furthermore, in one embodiment of the present invention, S1 includes: identifying thick, hard sandstone layers within a 120m range of the coal seam roof based on borehole columnar sections, and dividing them into subcritical layer I, subcritical layer II, and subcritical layer III; based on the fracture morphology of the subcritical layers, establishing an asymmetric T-shaped roof structure model for the initial pressure stage and a symmetric T-shaped roof structure model for the fully mined stage, respectively.

[0009] Furthermore, in one embodiment of the present invention, S2 further includes: calculating the additional static stress caused by mining formed by the failure of subcritical layers I, II, and III using an isosceles triangle distribution model; calculating the dynamic disturbance load generated by the failure of subcritical layers based on the theory of mine seismic energy release and the law of seismic wave propagation attenuation, and selecting S-wave as the main basis for calculating the dynamic disturbance load.

[0010] Furthermore, in one embodiment of the present invention, S3 further includes: dividing the difference between the total stress on the surrounding rock of the roadway and the roadway support stress by the critical stress for impact instability of the coal and rock mass to obtain the roadway surrounding rock impact instability index; judging the stability of the roadway surrounding rock based on the magnitude of the roadway surrounding rock impact instability index; determining that the roadway has experienced impact instability when the roadway surrounding rock impact instability index is greater than 1; determining that the roadway is in a critical instability state when the roadway surrounding rock impact instability index is equal to 1; and determining that the roadway has not experienced impact instability when the roadway surrounding rock impact instability index is less than 1.

[0011] To achieve the above objectives, another aspect of the present invention provides a device for determining the surrounding rock load and identifying its impact instability in coal mine adjacent working roadways, comprising: The roof structure modeling module is used to establish roof structure models for the initial pressure and full mining stages based on the borehole columnar diagrams near the adjacent working face, and to analyze the loading characteristics of the surrounding rock of the roadway at different mining stages. The total stress calculation module is used to calculate the lateral static load stress of the adjacent goaf, the advanced static load stress of the adjacent working face, and the dynamic load stress caused by the fracture of the overlying thick and hard rock layer on the roadway surrounding rock, so as to realize the total stress calculation under the superposition of the three loads. The impact instability index calculation module is used to calculate the impact instability index of the surrounding rock of the roadway based on the difference between the total stress on the roadway surrounding rock and the roadway support stress, combined with the critical stress of coal and rock mass impact instability, so as to determine whether the roadway has experienced impact instability. The prevention and control measures generation module is used to propose prevention and control measures for roadway rockburst based on the calculation results of the roadway surrounding rock impact instability index. These measures include using impact-resistant high-strength anchor bolts or constant resistance energy-absorbing anchor bolts for roadway support, or implementing hydraulic fracturing of the roof area and roof pre-fracture blasting for pressure relief in the advanced area of ​​the working face.

[0012] This invention discloses a method and apparatus for determining the load on the surrounding rock of a coal mine adjacent to a working face and for identifying its impact instability. By analyzing the loading state of the surrounding rock in this roadway, a method for determining the load on the surrounding rock is obtained. The stability of the surrounding rock is analyzed, and a criterion for identifying the impact instability of the surrounding rock in adjacent roadways is proposed, which is of great significance for the prevention and control of rockbursts in adjacent roadways. The method for determining the load on the surrounding rock of a coal mine adjacent to a working face and for identifying its impact instability can accurately determine the loading state of the surrounding rock at different mining stages, effectively identify its impact instability risk, and provide a scientific basis and methodological support for the prevention and control of rockbursts in deep coal mines.

[0013] Additional aspects and advantages of the invention will be set forth in part in the description which follows, and in part will be obvious from the description, or may be learned by practice of the invention. Attached Figure Description

[0014] The above and / or additional aspects and advantages of the present invention will become apparent and readily understood from the following description of the embodiments taken in conjunction with the accompanying drawings, wherein: Figure 1 This is a flowchart of a method for determining the surrounding rock load and identifying its impact instability in a coal mine adjacent roadway according to an embodiment of the present invention; Figure 2 This is a model diagram of the asymmetric T-shaped top plate structure of the working face during the initial pressure stage of the adjacent working face according to an embodiment of the present invention; Figure 3 This is a schematic diagram of the lateral static load distribution in the goaf area in front of the adjacent working face according to an embodiment of the present invention; Figure 4 This is a calculation model diagram of the additional stress caused by lateral mining in the goaf during the fully mined stage of working face II according to an embodiment of the present invention; Figure 5 This is a three-dimensional model of the roof of the working face during the initial pressure stage of the adjacent working face according to an embodiment of the present invention, and a schematic diagram of the superposition of three loads in the adjacent roadway. Figure 6 This is a schematic diagram of the superposition of dynamic and static loads in front of the working face during the initial pressure stage according to an embodiment of the present invention; Figure 7 This is a structural model diagram of a symmetrical T-shaped top plate of a working face in a periodic pressing stage according to an embodiment of the present invention. Figure 8 This is a model diagram of the symmetrical T-shaped roof structure of the working face during the fully activated stage of the adjacent working face according to an embodiment of the present invention; Figure 9 This is a three-dimensional model of the roof of the working face during the fully mined stage of the adjacent working face according to an embodiment of the present invention, and a schematic diagram of the superposition of three loads in the adjacent working face. Figure 10 This is a schematic diagram of the superposition of dynamic and static loads in front of the working face during the fully mobilized stage according to an embodiment of the present invention; Figure 11 This is a schematic diagram of a device for determining the load of surrounding rock in a coal mine adjacent roadway and for judging its impact instability, according to an embodiment of the present invention. Detailed Implementation

[0015] It should be noted that, unless otherwise specified, the embodiments and features described in the present invention can be combined with each other. The present invention will now be described in detail with reference to the accompanying drawings and embodiments.

[0016] To enable those skilled in the art to better understand the present invention, the technical solutions of the present invention will be clearly and completely described below with reference to the accompanying drawings of the embodiments of the present invention. Obviously, the described embodiments are only some embodiments of the present invention, and not all embodiments. Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort should fall within the scope of protection of the present invention.

[0017] The following describes, with reference to the accompanying drawings, a method and apparatus for determining the surrounding rock load and identifying its impact instability in a coal mine adjacent roadway, according to an embodiment of the present invention.

[0018] Figure 1 This is a flowchart of a method for determining the surrounding rock load and identifying impact instability in a coal mine adjacent roadway according to an embodiment of the present invention.

[0019] like Figure 1 As shown, the method includes, but is not limited to, the following steps: S1. Based on the borehole columnar section near the working face, a roof structure model is established for the initial pressure and full mining stages, and the loading characteristics of the surrounding rock in the roadway at different mining stages are analyzed. S2 calculates the lateral static stress of the adjacent goaf, the advanced static stress of the adjacent working face, and the dynamic stress caused by the fracture of the overlying thick and hard rock layer on the surrounding rock of the roadway, and realizes the total stress calculation under the superposition of the three loads. S3, based on the difference between the total stress on the surrounding rock of the roadway and the roadway support stress, combined with the critical stress of coal and rock mass impact instability, calculate the impact instability index of the surrounding rock of the roadway, which is used to determine whether the roadway has experienced impact instability. S4. Based on the calculation results of the rock impact instability index of the roadway surrounding rock, measures for preventing roadway rockburst are proposed, including using high-strength rockburst-resistant anchor bolts or constant resistance energy-absorbing anchor bolts for roadway support, or implementing hydraulic fracturing of the roof area and pre-fracturing blasting to relieve pressure in the advanced area of ​​the working face.

[0020] The method for determining the load on the surrounding rock of coal mine adjacent working roads and identifying its impact instability according to the embodiments of the present invention can accurately determine the loading state of the surrounding rock of adjacent working roads at different mining stages, effectively identify its impact instability risk, and provide scientific basis and methodological support for the prevention and control of rockbursts in deep coal mines.

[0021] The following describes in detail, with reference to the accompanying drawings, a method for determining the surrounding rock load and identifying its impact instability in a coal mine adjacent roadway according to an embodiment of the present invention.

[0022] This invention analyzes the loading state of the surrounding rock in coal mine roadways near working holes, obtains a method for determining the load on the surrounding rock, analyzes the stability of the surrounding rock, and proposes a criterion for rockburst instability of the surrounding rock in roadways near working holes, which is of great significance for the prevention and control of rockbursts in such roadways. The invention elucidates the method for determining the load on the surrounding rock in coal mine roadways near working holes and the method for identifying rockburst instability of the surrounding rock, mainly including the following: Based on the lithology revealed by a typical borehole column near the working face of a rockburst mine, a roof structure model of the working face during the initial pressure and full mining stages, a three-dimensional roof model of the working face during the initial pressure and full mining stages, and a schematic diagram of the superposition of three loads on the working face roadway were established. The loading state of the working face roadway and the load transfer law of roof failure at different mining stages were analyzed, a method for determining the surrounding rock load of the working face roadway was obtained, and finally, a criterion for the rockburst instability of the surrounding rock of the working face roadway was proposed.

[0023] 1. Calculation of roof structure and surrounding rock stress in adjacent working tunnels during the initial pressure stage: Two working faces (working faces I and II) have been mined in a certain mining area of ​​a mine. Based on the overlying strata revealed by typical borehole columns near the working faces, key strata were identified. Analysis showed that the sandstone strata with greater strength and thickness within 120m of the coal seam roof were identified as sub-key strata I, II, and III. The fracture structure morphology of the near-field key strata significantly affects the stress in the surrounding rock of the roadway. Therefore, the load-bearing state of the surrounding rock should be analyzed based on the fracture structure of the key strata to obtain a method for calculating the stress in the surrounding rock, laying the foundation for the analysis of the roadway rockburst mechanism.

[0024] When working face II has been mined, and the adjacent working face has reached the initial pressure stage, the roof of the goaf in the working face forms an asymmetrical T-shaped structure, such as... Figure 2 As shown, subcritical layers I, II, and III overlying working face II are all fractured, forming a "multi-layered inclined masonry beam" structure along the working face layout direction. Subcritical layer I overlying the adjacent working face is fractured, while subcritical layers II and III remain intact, resulting in a cantilever roof phenomenon and forming a "single-layered inclined masonry beam" structure along the working face layout direction. At this time, the mine pressure in the adjacent working face is relatively mild, and the roof load is transferred to the coal and rock mass at both ends of the working face, leading to stress concentration and energy accumulation in localized areas, providing conditions for inducing roadway rockbursts.

[0025] At different mining stages of the adjacent working face, the lateral static load stress on the adjacent roadway ahead of the working face II goaf is basically the same, and the distribution pattern is shown in the figure below. Figure 3 As shown.

[0026] The lateral static load stress on the adjacent roadway in the goaf area of ​​working face II includes the foundation static load stress and the additional static load stress caused by mining: (1) In the formula, s j The lateral static load stress of the goaf in the adjacent roadway during the full mining stage of working face II; s y The static load stress of the foundation is caused by the self-weight of the overlying rock strata; ∆ s I Additional static stress caused by mining in subcritical layer I; ∆ s II Additional static stress caused by mining in subcritical layer II; ∆ s III The mining-induced additional static stress is generated in the subcritical layer III.

[0027] The additional static load stress experienced by the adjacent goaf roadway is mainly formed by the static load transferred from the fractures of subcritical layers I, II, and III. The fractured rock blocks of subcritical layers I, II, and III transfer the static load to the goaf and the end of the working face, which is approximately equal to half of the self-weight of the fractured rock blocks of subcritical layers I, II, and III and the load transferred by the overlying rock strata.

[0028] Based on on-site monitoring data and engineering experience, it is assumed that the additional static stress caused by mining is transferred to the lateral coal and rock mass of the goaf in an isosceles triangular distribution. The origin of the coordinate system is taken as the end of working face II. O , with the vertical direction as s The axis is the horizontal direction. x Axis, establish a calculation model for the additional stress caused by lateral mining in the goaf of working face II, such as Figure 4 As shown.

[0029] The lateral static stress in the goaf mainly includes the foundation static stress (self-weight stress) from the self-weight of the roof strata. sy and the additional mining static stress ∆ transmitted from the overlying strata caused by mining. s .

[0030] Due to coal seam mining, the static load stress on the foundation acting on the lateral coal and rock mass of the goaf increases. s y It varies significantly with increasing distance from the coal seam. Figure 3 middle OF Within range s y linearly increasing from 0 to cH ,exist F Click on the left s y equal cH static stress of foundation s y It can be represented by a piecewise function.

[0031] (2) In the formula, c Unit weight of the rock strata; H The depth of the adjacent working face; α The angle of rock strata movement can be approximated as 70°. h 0~ h 6 represents the thickness of each rock layer above the coal seam.

[0032] based on Figure 3 Mesosceles triangle AOB The geometric relationship, during the fully developed mining stage of working face II, due to the fracture of subcritical layer I, the additional static stress ∆ formed on the lateral coal and rock mass of the goaf. s I It can be calculated using the following formula.

[0033] (3) In the formula, s Imax The peak value of additional static load stress caused by mining during the fully mining stage of working face II, specifically the subcritical layer I and its overlying strata.

[0034] As analyzed above, half of the weight of the fractured rock block in subcritical layer I and its overlying strata is transferred to the end of the working face. Figure 3 Additional static stress ∆ in mining s I The distribution pattern can be obtained s Imax The calculation formula is as follows: (4) In the formula, Q zIThe self-weight of the fractured rock block of sub-critical layer I and its overlying rock strata during the full mining stage of working face II; c i This refers to the unit weight of the corresponding rock stratum; l 1 represents the length of the fracture block in the subcritical layer I.

[0035] The calculation of dynamic additional static stress ∆ is adopted. σI The same method, based on Figure 3 Mesosceles triangle COD The geometric relationship. During the fully developed mining stage of working face II, the additional static stress ∆ caused by mining on the lateral coal and rock mass of the goaf due to the subcritical layer II and its corresponding overlying strata. σII It can be calculated using the following formula.

[0036] (5) In the formula, s IImax The peak value of additional static load stress caused by mining during the fully mining stage of working face II, specifically the subcritical layer II and its overlying strata.

[0037] As analyzed above, half of the self-weight of the fractured rock block in subcritical layer II and its overlying strata is transferred to the end of the working face. The peak static stress caused by mining is calculated using this method. s Imax The same method, according to Figure 3 Additional static stress ∆ in mining s II The distribution pattern can be obtained s IImax The calculation formula is as follows: (6) In the formula, Q zII The self-weight of the fractured rock block of the sub-key layer II and its overlying rock strata during the full mining stage of working face II; l 2 represents the length of the fracture block in the subcritical layer II.

[0038] Similarly, based on Figure 3 Mesosceles triangle EOF The geometric relationship, during the fully developed mining stage of working face II, due to the load transferred by sub-critical layer III and its corresponding overlying strata, the additional static stress ∆ formed on the lateral coal and rock mass of the goaf is due to mining. s III It can be calculated using the following formula.

[0039] (7) In the formula, s IIImax The peak value of additional static load stress caused by mining during the fully mining stage of working face II, specifically the subcritical layer III and its overlying strata.

[0040] As analyzed above, half of the load transferred from subcritical layer III and its corresponding overlying strata is transmitted to the working face end. The peak static stress caused by mining is calculated using this method. s Imax Using the same method, based on the additional static load stress ∆ in the mining operation. s III The distribution pattern can be obtained s IIImax The calculation formula is as follows: (8) In the formula, Q zIII The load is transferred to the fractured rock block of the sub-critical layer III and its corresponding overlying strata during the full mining stage of working face II. c 7 represents the unit weight of the overlying load layer; h 7 represents the thickness of the overlying load layer; K z The load transfer coefficient of the overlying strata; l 3 represents the length of the fracture block in the subcritical layer III.

[0041] The lateral static stress of the surrounding rock of the adjacent roadway in the working face II goaf area can be calculated using equations (1) to (8). s j .

[0042] The surrounding rock of the adjacent roadway ahead of the working face is mainly affected by the superposition of the lateral support pressure of the goaf area of ​​working face II, the advance support pressure of this working face, and the dynamic load of the fracture of the overlying thick and hard rock strata. When the total load reaches the critical load for coal and rock impact instability, it may induce rockburst. The load on the coal and rock mass ahead of the working face varies in different mining stages. It is necessary to analyze the roof structure specifically according to each mining stage, and then describe the loading state of the adjacent roadway in detail.

[0043] Three-dimensional model of the roof of the working face during the initial pressure stage and lateral static stress of the goaf area of ​​working face II. s j Pre-load static stress of adjacent working face s jc , breaking dynamic load of overlying thick hard rock layer s dc Distribution pattern diagram as follows Figure 5 As shown.

[0044] The top plate structure and the superposition of dynamic and static loads in front of the working face along the working face advancement direction from the initial cutting point to the initial pressure stage, such as... Figure 6 As shown.

[0045] During the initial pressure stage of the adjacent working face, the static load on the surrounding rock near the adjacent end of the working face and in the adjacent roadway in front of the working face is mainly the lateral static stress σ formed by the goaf area of ​​working face II. j and the advanced static stress formed by the advanced support pressure of the adjacent working face s jc .

[0046] The magnitude and distribution of the support pressure at the working face are influenced by many factors, such as the mining depth, mining height, roof lithology, and structure. The limit equilibrium theory is typically used to analyze the distribution of the roof support pressure at the working face, yielding the support pressure distribution in the limit equilibrium zone and the elastic zone as follows: (9) in, , ; In the formula, s a It is vertical stress; d It is the triaxial stress coefficient; P For working face support resistance; c , f The cohesion and internal friction angle of the coal; f The coefficient of friction between the coal seam and the roof and floor surfaces; m To raise the working face; x The distance from any point to the coal face; x 0 represents the width of the limit equilibrium zone; K ´ represents the stress concentration factor; c Unit weight of the rock strata; H The depth of the working face; l This is the lateral pressure coefficient.

[0047] The advanced static load stresses σjc and σjz of the initial pressure and periodic pressure (full mining) stages of the adjacent working face can be calculated from equation (9).

[0048] During the initial pressure stage of the adjacent working face, the main source of dynamic load on the surrounding rock of the adjacent roadway is the disturbance dynamic load generated when the subcritical stratum I first fractures. The fracture of subcritical stratum I generates mine tremors, which release stress waves that propagate laterally to the coal and rock mass in the goaf area. Figure 6 The magnitude of the dynamic load resulting from the disturbance is related to factors such as the energy level at the earthquake source, the propagation mode, and the attenuation characteristics.

[0049] The seismic energy obtained from on-site microseismic monitoring is generally η times the energy released by roof failure. Combining this with the formula for calculating the accumulated elastic energy before the failure of key layer I, the seismic energy generated by the failure of sub-key layer I during the initial pressure stage can be obtained as follows: (10) In the formula, U kFor mine seismic energy; or This is the energy conversion coefficient, which can be taken as 0.1% to 1%. U wzⅠ This refers to the elastic energy accumulated before the subcritical layer I fractures. q 1 represents the load of the overlying strata of subcritical layer I; E The elastic modulus of the top stratum; J Let be the moment of inertia.

[0050] The literature uses in-situ test data to fit and analyze the propagation law of seismic waves in underground coal mines, and obtains the following relationship between peak particle vibration velocity and seismic energy: (11) In the formula, V pm denoted as peak vibration velocity of the particle.

[0051] Considering the attenuation phenomenon of seismic waves during propagation, the literature conducted in-situ experiments on the propagation law of seismic waves in coal mines, and obtained the attenuation law of particle vibration velocity caused by mine seismic activity as follows: (12) In the formula, V r For the distance of transmission r Peak vibrational velocity of a particle at that point; V 0 represents the peak vibration velocity of the particles at the earthquake source; r 1 represents the distance the vibration wave travels; β 1 represents the attenuation index of the seismic wave propagation caused by the failure of the subcritical layer I, which is taken as 1.526.

[0052] The dynamic load generated by seismic stress waves in coal and rock mass can be expressed as: (13) In the formula, s dP / S These are the dynamic loads generated by P-wave and S-wave, respectively. r The density of the medium; V P / S These are the propagation velocities of P-wave and S-wave, respectively; V P ) P / S These are the peak vibration velocities of the particles caused by the propagation of P-waves and S-waves, respectively.

[0053] Studies have shown that the amplitude of S-waves in high-energy mine seismic events caused by the failure of key overburden layers is much larger than that of P-waves, and the dynamic damage intensity is also greater. Therefore, S-waves are selected for calculating the dynamic load of roof failure disturbance.

[0054] From equations (10) to (13), the dynamic disturbance load generated when the subcritical layer I fails during the initial pressure stage of the adjacent working face is: (14) Based on the above analysis, the total stress on the surrounding rock of the adjacent working face during the initial pressure stage is: (15) 2. Calculation of roof structure and surrounding rock stress in the fully mined stage: After the initial pressure of the adjacent working face, as the working face continues to advance, it will successively undergo cyclic pressure and full mining stages. During the cyclic pressure stage of the adjacent working face, the overlying subcritical layers I, II, and III all fracture, forming a "multi-layered inclined masonry beam" structure along the working face's layout direction. The roof of the goaf area forms a symmetrical T-shaped structure, such as... Figure 7 As shown, the failure of subcritical layers II and III will exert dynamic loads on subcritical layer I, the working face, and the coal and rock mass at both ends. If the loaded coal and rock mass already has high stress concentration or a large amount of elastic energy accumulation, rockburst in adjacent roadways is likely to occur under dynamic loads.

[0055] During the fully developed stage of the adjacent working face, the roof of the goaf area forms a symmetrical T-shaped structure, such as... Figure 8 As shown. At this time, the high-level, thick, hard rock strata on the working face undergo large-scale movement, fully releasing the elastic energy stored in the roof. This increases the dynamic load disturbance of the subcritical strata I, II, and III on the coal and rock mass below, forming a strong dynamic load disturbance. When the total stress or energy on the coal and rock mass exceeds its critical value for impact instability, rockburst occurs in the adjacent roadway. Compared with the periodic pressure stage, due to the large-scale movement of the overlying rock strata on the working face, a more intense dynamic load effect is formed, and the probability of rockburst in the adjacent roadway is greater in this stage.

[0056] Three-dimensional model of the roof of the working face during the fully developed stage of the adjacent working face and lateral static stress of the goaf area of ​​working face II. s j Pre-load static stress of adjacent working face s jz , breaking dynamic load of overlying thick hard rock layer s dz Distribution pattern diagram as follows Figure 9 As shown.

[0057] The above analysis shows that rockbursts are more likely to occur in adjacent roadways during the fully mined-out stage than during the periodic pressure stage. Therefore, after the initial pressure on the working face, the analysis mainly focuses on the load state of the adjacent roadways ahead of the working face during the fully mined-out stage. This includes the roof structure along the working face advance direction and the superposition of dynamic and static loads ahead of the working face when the adjacent working face reaches the fully mined-out stage. Figure 10 As shown.

[0058] At this point, the static load source of the surrounding rock in the adjacent working face is basically the same as that in the initial pressure stage. The main source of the dynamic load is the disturbance dynamic load generated when the subcritical layers I, II, and III fail in cycles. Figure 10 When the three subcritical layers I, II, and III break simultaneously, the resulting disturbance dynamic load is the largest. The stress calculation of the surrounding rock in the adjacent roadway should consider the worst stress environment and should be determined according to the strong disturbance dynamic load generated when the three subcritical layers break simultaneously.

[0059] Similarly, the disturbance dynamic load generated when the subcritical layers I, II, and III simultaneously break during the fully mined stage of the adjacent working face is: (16) In the formula, r zⅠ , r zⅠⅠ , r zⅠⅠⅠ The distances of seismic wave propagation caused by the failure of subcritical layers I, II, and III are respectively. q zⅠ , q zⅠⅠ , q zⅠⅠⅠ The loads of the overlying strata of subcritical layers I, II, and III are respectively; l zⅠ , l zⅠⅠ , l zⅠⅠⅠ These are the fracture step distances for subcritical layers I, II, and III, respectively. β 2. β 3 represents the attenuation index of seismic wave propagation caused by the failure of subcritical layers II and III.

[0060] Based on the above analysis, the total stress on the surrounding rock of the adjacent working face during the fully activated stage of mining can be obtained as follows: (17) 3. Stability analysis of surrounding rock in adjacent tunnels: The critical stress required for impact instability of coal and rock mass is definite under specific coal seam occurrence conditions and mining technology. It is mainly influenced by factors such as mining depth, coal and rock lithology, impact tendency, mining area layout, and mining technology. Assuming the critical stress for impact instability of coal and rock mass under certain mining conditions is... s bmin .

[0061] According to equations (15) and (17), the total stress experienced by the adjacent roadway during the initial pressure and full mining stages of the adjacent working face is: s zc , szz Define the total stress on the surrounding rock of the tunnel. s zc ( s zz ) and roadway support stress ψ The difference between the values ​​is related to the critical stress for impact instability of coal and rock mass. s bmin The ratio is the impact instability index of the surrounding rock in the tunnel. K b The stability of the surrounding rock in the tunnel can be categorized into three cases, which can be represented by equation (18): (1) When the surrounding rock of the roadway is impacted and becomes unstable K b When the value is greater than 1, the roadway will experience shock instability.

[0062] (2) When the surrounding rock of the roadway is impacted and becomes unstable K b When the value equals 1, the roadway is in a critical state of instability.

[0063] (3) When the surrounding rock of the roadway is impacted and becomes unstable K b When the value is less than 1, the roadway will not experience shock instability.

[0064] (18) Based on the above analysis, given a relatively fixed critical stress σbmin for coal and rock mass impact instability, to improve the stability of the surrounding rock in roadways and mitigate or avoid rockbursts, measures should be taken to reduce the total stress on the surrounding rock or to increase the strength of the surrounding rock support, such as using high-strength impact-resistant anchor bolts or constant-resistance energy-absorbing anchor bolts for strong support. Regarding the total stress on the surrounding rock, the most significant factor is the dynamic load generated when the critical overburden layer at the working face fractures. Therefore, effective stress relief measures are needed to reduce the dynamic load formed by the fracture of the thick, hard overburden at the working face, thereby reducing the risk of roadway impact.

[0065] The method for determining the load on the surrounding rock of coal mine adjacent working roads and identifying its impact instability according to embodiments of the present invention can accurately determine the loading state of the surrounding rock of adjacent working roads at different mining stages, effectively identify its impact instability risk, and provide scientific basis and methodological support for the prevention and control of rockbursts in deep coal mines.

[0066] To achieve the above embodiments, such as Figure 11 As shown, this embodiment also provides a device 10 for determining the load of the surrounding rock in a coal mine adjacent roadway and for judging its impact instability. The device 10 includes: a roof structure modeling module 100, a total stress calculation module 200, an impact instability index calculation module 300, and a prevention and control measure generation module 400.

[0067] The roof structure modeling module 100 is used to establish roof structure models for the initial pressure and full mining stages based on the borehole columnar diagram near the adjacent working face, and to analyze the loading characteristics of the surrounding rock of the roadway at different mining stages. The total stress calculation module 200 is used to calculate the lateral static load stress of the adjacent goaf, the advanced static load stress of the adjacent working face, and the dynamic load stress caused by the fracture of the overlying thick and hard rock layer on the surrounding rock of the roadway, so as to realize the total stress calculation under the superposition of the three loads. The Impact Instability Index Calculation Module 300 is used to calculate the impact instability index of the surrounding rock of the roadway based on the difference between the total stress on the roadway surrounding rock and the roadway support stress, combined with the critical stress of coal and rock mass impact instability, in order to determine whether the roadway has experienced impact instability. The prevention and control measures generation module 400 is used to propose prevention and control measures for roadway rockburst based on the calculation results of the roadway surrounding rock impact instability index. These measures include using impact-resistant high-strength anchor bolts or constant resistance energy-absorbing anchor bolts for roadway support, or implementing hydraulic fracturing of the roof area and roof pre-fracture blasting for pressure relief in the advanced area of ​​the working face.

[0068] Furthermore, the aforementioned top slab structure modeling module 100 is also used for: Specifically, based on the borehole columnar section, the thick and hard sandstone layer within 120m of the coal seam roof was identified and divided into subcritical layer I, subcritical layer II and subcritical layer III. Based on the fracture morphology of the subcritical layers, an asymmetric T-shaped roof structure model for the initial pressure stage and a symmetric T-shaped roof structure model for the fully mined stage were established respectively.

[0069] Furthermore, the total stress calculation module 200 is also used for: Specifically, an isosceles triangle distribution model was used to calculate the additional static stress caused by the failure of subcritical layers I, II, and III. Based on the theory of mine seismic energy release and the law of seismic wave propagation attenuation, the dynamic disturbance load generated by the failure of subcritical layers was calculated, and the S-wave was selected as the main basis for calculating the dynamic disturbance load.

[0070] Furthermore, the shock instability index calculation module 300 is also used for: Specifically, the difference between the total stress on the surrounding rock of the roadway and the roadway support stress is divided by the critical stress for impact instability of the coal and rock mass to obtain the roadway surrounding rock impact instability index. The stability of the roadway surrounding rock is judged according to the magnitude of the roadway surrounding rock impact instability index. When the roadway surrounding rock impact instability index is greater than 1, the roadway is determined to have experienced impact instability. When the roadway surrounding rock impact instability index is equal to 1, the roadway is determined to be in a critical instability state. When the roadway surrounding rock impact instability index is less than 1, the roadway is determined not to have experienced impact instability.

[0071] According to an embodiment of the present invention, a method and apparatus for determining the load of surrounding rock in a coal mine adjacent to a working hole and for judging its impact instability are provided. By analyzing the loading state of the surrounding rock in the coal mine adjacent to a working hole, a method for determining the load of the surrounding rock is obtained, the stability of the surrounding rock is analyzed, and a criterion for judging the impact instability of the surrounding rock in the adjacent to a working hole is proposed, which is of great significance for the prevention and control of rockburst in adjacent to a working hole.

[0072] In the description of this specification, the references to terms such as "one embodiment," "some embodiments," "example," "specific example," or "some examples," etc., indicate that a specific feature, structure, material, or characteristic described in connection with that embodiment or example is included in at least one embodiment or example of the present invention. In this specification, the illustrative expressions of the above terms do not necessarily refer to the same embodiment or example. Furthermore, the specific features, structures, materials, or characteristics described may be combined in any suitable manner in one or more embodiments or examples. Moreover, without contradiction, those skilled in the art can combine and integrate the different embodiments or examples described in this specification, as well as the features of different embodiments or examples.

[0073] Furthermore, the terms "first" and "second" are used for descriptive purposes only and should not be construed as indicating or implying relative importance or implicitly specifying the number of technical features indicated. Thus, a feature defined as "first" or "second" may explicitly or implicitly include at least one of that feature. In the description of this invention, "a plurality of" means at least two, such as two, three, etc., unless otherwise explicitly specified.

Claims

1. A method for determining the surrounding rock load of a coal mine adjacent empty roadway and its impact instability discrimination, characterized in that, The method comprises the following steps: S1, establishing a roof structure model of a primary pressure and a fully-mined stage based on a borehole column chart near an adjacent goaf working face, and analyzing load bearing characteristics of a roadway surrounding rock in different mining stages; S2, calculating lateral static load stress of the roadway surrounding rock from an adjacent goaf, advanced static load stress of the roadway surrounding rock from the adjacent goaf working face, and dynamic load stress caused by breaking of an overlying thick hard rock layer, and realizing total stress calculation under three load superposition; S3, calculating a rock burst instability index of the roadway surrounding rock according to a difference between total stress of the roadway surrounding rock and roadway support stress, and combining a coal rock mass rock burst instability critical stress, and judging whether the roadway has rock burst instability; S4, proposing rock burst prevention and control measures for the roadway according to a calculation result of the rock burst instability index of the roadway surrounding rock, including roadway support by using impact-resistant high-strength anchor rods or constant-resistance energy-absorbing anchor rods, or implementing roof regional hydraulic fracturing or roof pre-splitting blasting pressure relief treatment in an advanced area of the working face.

2. The method of claim 1, wherein, The S1 comprises: According to the borehole column chart, key layers of rock layers within a range of 120m of the coal seam roof are distinguished, and the rock layers are divided into sub-key layer I, sub-key layer II and sub-key layer III; Based on the breaking form of the sub-key layer, an asymmetric T-shaped roof structure model of a primary pressure stage and a symmetric T-shaped roof structure model of a fully-mined stage are respectively established.

3. The method of claim 1, wherein, The S2 further comprises: The mining additional static load stress formed by breaking of the sub-key layers I, II and III is calculated by using an isosceles triangle distribution model; Based on a mine earthquake energy release theory and a vibration wave propagation attenuation law, a disturbance dynamic load generated by breaking of the sub-key layer is calculated, and S wave is selected as a main disturbance dynamic load calculation basis.

4. The method of claim 1, wherein, The S3 further comprises: The rock burst instability index of the roadway surrounding rock is obtained by dividing the difference between the total stress of the roadway surrounding rock and the roadway support stress by the coal rock mass rock burst instability critical stress; The stability of the roadway surrounding rock is judged according to the size of the rock burst instability index of the roadway surrounding rock, when the rock burst instability index of the roadway surrounding rock is greater than 1, it is determined that the roadway has rock burst instability, when the rock burst instability index of the roadway surrounding rock is equal to 1, it is determined that the roadway is in a critical instability state, and when the rock burst instability index of the roadway surrounding rock is less than 1, it is determined that the roadway does not have rock burst instability.

5. A device for determining the load of surrounding rock of a coal mine adjacent to an empty roadway and distinguishing the impact instability thereof, characterized in that, It comprises: A roof structure modeling module is used to establish a roof structure model of a primary pressure and a fully-mined stage based on a borehole column chart near an adjacent goaf working face, and analyze load bearing characteristics of a roadway surrounding rock in different mining stages; A total stress calculation module is used to calculate lateral static load stress of the roadway surrounding rock from an adjacent goaf, advanced static load stress of the roadway surrounding rock from the adjacent goaf working face, and dynamic load stress caused by breaking of an overlying thick hard rock layer, and realize total stress calculation under three load superposition; A rock burst instability index calculation module is used to calculate a rock burst instability index of the roadway surrounding rock according to a difference between total stress of the roadway surrounding rock and roadway support stress, and combine a coal rock mass rock burst instability critical stress, and judge whether the roadway has rock burst instability; A prevention and control measure generation module is used to propose rock burst prevention and control measures for the roadway according to a calculation result of the rock burst instability index of the roadway surrounding rock, including roadway support by using impact-resistant high-strength anchor rods or constant-resistance energy-absorbing anchor rods, or implementing roof regional hydraulic fracturing or roof pre-splitting blasting pressure relief treatment in an advanced area of the working face.

6. The apparatus of claim 5, wherein, The roof structure modeling module is further used to Based on the borehole columnar section, the rock strata within a 120m range of the coal seam roof were identified as key strata and divided into sub-key strata I, sub-key strata II and sub-key strata III. Based on the fracture morphology of the subcritical layer, an asymmetric T-shaped roof structure model for the initial pressure stage and a symmetric T-shaped roof structure model for the fully mined stage were established respectively.

7. The apparatus of claim 5, wherein, The total stress calculation module is also used for: The additional static stress caused by mining is calculated by using an isosceles triangle distribution model to calculate the stress caused by the failure of subcritical layers I, II, and III. Based on the theory of mine seismic energy release and the law of seismic wave propagation attenuation, the dynamic load of disturbance generated by the failure of the subcritical layer is calculated, and the S-wave is selected as the main basis for the calculation of dynamic load of disturbance.

8. The apparatus of claim 5, wherein, The shock instability index calculation module is also used for: The difference between the total stress on the surrounding rock of the roadway and the roadway support stress is divided by the critical stress for impact instability of the coal and rock mass to obtain the impact instability index of the surrounding rock of the roadway. The stability of the surrounding rock of a roadway is determined by the magnitude of its impact instability index. When the impact instability index is greater than 1, the roadway is considered to have experienced impact instability. When the impact instability index is equal to 1, the roadway is considered to be in a critical instability state. When the impact instability index is less than 1, the roadway is considered not to have experienced impact instability.

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