Method for reinforcing mined-out side concrete wall of coal pillar in mining section

Abstract

The invention discloses a mining section coal pillar goaf side concrete wall reinforcing method, and relates to the field of coal mining, and the method comprises the following steps: collecting section coal pillar geological parameters; calculating a boundary position of a plastic zone according to geological parameters of the coal pillar, and dividing the coal pillar into a broken zone, a fractured zone and an original structure zone; basic roof strata are determined according to geological parameters of the coal pillar, the bearing capacity of the basic roof strata is calculated, the basic roof strata with the minimum bearing capacity serves as a roof cutting target layer, and the hydrofracture roof cutting drilling depth needed for penetrating through the target layer is determined; the broken area range serves as the influence range of surrounding rock pressure calculation, the static load q applied to the concrete wall by the surrounding rock pressure is calculated, the static load q is corrected through a dynamic load coefficient, and a supporting load Q is obtained; and according to the support load Q, reinforcement parameters are determined. Aiming at the surrounding rock deformation and instability of the adjacent roadway on the section coal pillar gob side under the deep well mining condition, the method effectively controls the surrounding rock deformation and instability of the adjacent roadway caused by working face mining disturbance.

Description

Technical Field

[0001] The present application relates to the field of coal mining, and in particular to a method for reinforcing a concrete wall on the goaf side of a coal pillar in a mining section. Background Art

[0002] With the continuous deepening of coal mining in my country, shallow coal resources are becoming increasingly depleted, and coal mining is gradually shifting to deeper areas. Deep-shaft mining has become an inevitable trend in the development of my country's coal industry, with mining depths generally reaching 600-800 meters or even deeper. Under deep-shaft mining conditions, problems such as high ground stress, complex rock properties, and strong mining disturbances are becoming increasingly prominent, posing severe challenges to the safe and efficient operation of coal mines. In particular, in deep-shaft mining areas that utilize high-cut height and longwall mining techniques, the disturbance effect of working face mining on adjacent roadways is even more significant. Controlling roadway surrounding rock deformation has become a key technical challenge restricting safe production in deep-shaft coal mines.

[0003] During deep-well mining, the segmental coal pillar, as the coal body between the goaf formed by working face mining and the area to be mined, plays an important role in protecting the adjacent roadways. However, due to the high ground stress and relatively low coal body strength in deep wells, the segmental coal pillar is subjected to strong supporting pressure after the working face is mined, and is prone to plastic deformation or even destruction, resulting in a significant decrease in its bearing capacity. At the same time, the roof rock layer in deep-well mining is often of low strength and easily weathered and broken, the roadway design section is large, and the surrounding rock's self-bearing capacity is insufficient. Under geological conditions with developed structures and poor rock integrity, traditional anchor cable support is difficult to establish an effective bearing structure in the loose and broken surrounding rock. The roadway surrounding rock exhibits phenomena such as flow extrusion and net bag slag falling, and the deformation and instability of the surrounding rock becomes increasingly serious.

[0004] Existing coal pillar support and roadway reinforcement technologies primarily rely on passive measures such as increasing support density and improving support strength. These measures lack a scientific understanding of the distribution patterns of the coal pillar's plastic zone and accurate calculation of surrounding rock loads. Support design is often determined empirically and lacks a theoretical basis. Furthermore, top-cutting and roadway reinforcement technologies are typically applied independently, lacking collaborative optimization design and making it difficult to form an effective mechanical coupling and coordinated load-bearing system. Therefore, there is an urgent need to establish a scientific method for reinforcing the goaf-side concrete walls of coal pillars in deep-mine sections. Through precise analysis of the coal pillar's plastic zone, reasonable top-cutting and roadway reinforcement design, and scientific determination of concrete wall support parameters, mechanical coupling and coordinated load-bearing between top-cutting and roadway reinforcement can be achieved, effectively controlling the deformation and instability of the surrounding rock of adjacent roadways under deep-mine conditions.

[0005] For example, the related art CN118351952A discloses a method for calculating the stability of a segmented coal pillar, including the following steps: S1. Calculation of the bending elastic energy of the basic roof on the air side: Establish a mechanical model of the basic roof on the air side, and calculate the bending elastic energy UW of the basic roof on the air side based on the mechanical model; S2. Calculation of the yield elastic energy of the segmented coal pillar: Establish a lateral stress model of the segmented coal pillar, obtain a calculation formula for the lateral stress distribution of the segmented coal pillar, and obtain the distribution expression of the yield elastic energy Es of the segmented coal pillar elastic zone along the width of the coal pillar based on the calculation formula for the lateral stress distribution of the segmented coal pillar; S3. Using UW>Es as the criterion for coal body instability in the elastic zone of the segmented coal pillar, analyze and judge the stability of the segmented coal pillar. However, this application adopts the method of overall elastic energy comparison, treating the coal pillar as a homogeneous body for analysis, and ignores the complexity of the stress distribution within the coal pillar. Summary of the Invention

[0006] In response to the deformation and instability of the surrounding rock of the adjacent tunnels on the side of the coal pillar goaf in the section under deep well mining conditions, this application provides a method for reinforcing the concrete wall on the side of the coal pillar goaf in the mining section. Through the mechanical coupling of the top cutting and the concrete wall reinforcement, the deformation and instability of the surrounding rock caused by the working face mining disturbance to the adjacent tunnels can be effectively controlled.

[0007] The present application provides a method for reinforcing the concrete wall of the goaf side of the coal pillar in the mining section, comprising: collecting geological parameters of the coal pillar in the section, wherein the coal pillar in the section is the coal body between the goaf formed by mining of the working face and the area to be mined; calculating the boundary position of the plastic zone according to the geological parameters of the coal pillar; , according to the position of the plastic zone boundary The coal pillar of the section is divided into the crushed zone, the fractured zone and the original structure zone; among them, the boundary position of the plastic zone It refers to the horizontal distance from the edge of the goaf to the critical position where the coal body changes from plastic deformation state to elastic state. This position is the dividing line of the stress state of the coal pillar. The bearing pressure at the position reaches its maximum value, and the coal body beyond this position still maintains an elastic state and complete bearing capacity.

[0008] According to the geological parameters of the coal pillar, the basic top rock layer is determined, the bearing capacity of each basic top rock layer is calculated, the basic top rock layer with the smallest bearing capacity is used as the top cutting target layer, and the hydraulic fracturing top cutting drilling depth required to penetrate the target layer is determined; the crushing zone range is used as the influence range of the surrounding rock pressure calculation, the static load q imposed by the surrounding rock pressure on the concrete wall is calculated, and the dynamic load coefficient is used to correct the static load q to obtain the support load Q borne by the concrete wall; the dynamic load coefficient is used to correct the static load to obtain the support load Q; according to the support load Q, the concrete strength grade of the concrete wall, the thickness of the concrete wall in the direction perpendicular to the tunnel axis in the tunnel section, and the horizontal distance between the top of the concrete wall and the edge of the side roof of the goaf are determined.

[0009] Specifically, based on the theory of elastic-plastic mechanics and the Mohr-Coulomb failure criterion, the coal pillar will undergo elastic-plastic deformation under the action of support pressure after mining at the working face. By establishing a vertical stress distribution model in the plastic zone and combining the boundary conditions to determine the range of the plastic zone, it is essentially a quantitative description of the stress-strain relationship of the coal pillar. The boundary position of the plastic zone The calculation reveals the critical position where the coal pillar changes from elastic state to plastic state. The three-zone division based on this scientifically reflects the differences in bearing state and stability of different parts of the coal pillar.

[0010] Furthermore, the geological parameters of the coal pillar include: uniaxial compressive strength of the coal seam , the average bulk density γ of the overlying rock layer, the buried depth H, the coal seam mining thickness h, the coal seam compression angle α, the cohesion C between the coal seam and the roof and floor, as well as the thickness of each rock layer and the rock layer expansion coefficient .

[0011] Among them, the uniaxial compressive strength of coal seam This refers to the maximum compressive stress a coal sample can withstand under uniaxial compression until it reaches failure, measured in MPa. This parameter reflects the inherent strength characteristics of the coal mass and is a key mechanical indicator for calculating the boundary position and bearing capacity of the coal pillar's plastic zone. It directly affects the deformation and failure mode of the coal pillar under abutment pressure.

[0012] The average bulk density γ of the overburden is the average unit volume weight of all rock layers above the coal seam, expressed in kN / m³. This parameter is used to calculate the vertical stress exerted by the overburden on the coal pillar. It is a fundamental parameter for determining the loading state and plastic zone distribution of the coal pillar, and directly affects the magnitude and distribution of the abutment pressure.

[0013] Burial depth, H, is the vertical distance from the ground surface to the coal seam roof, measured in meters. Burial depth determines the stress borne by the overlying rock strata under the coal seam's own weight and is an important boundary condition for calculating the initial stress state of the coal pillar and stress redistribution after mining.

[0014] Coal seam mining thickness h refers to the actual thickness of the coal seam mined at the working face, measured in meters. This parameter affects the geometric dimensions of the goaf and the collapse characteristics of the overlying strata. It is an important geometric parameter for calculating the extent of the coal pillar's plastic zone and the concrete wall support load.

[0015] The coal seam compression angle α is the angle between the compressive stress and the normal to the coal seam plane during compression and deformation, measured in degrees. This parameter reflects the compression and deformation characteristics of the coal seam and is used to establish a vertical stress distribution model in the plastic zone, influencing the calculation results of the plastic zone boundary position of the coal pillar.

[0016] Cohesion C between the coal seam and the roof and floor plates refers to the bond strength between the coal seam and the adjacent rock strata above and below it, measured in MPa. This parameter reflects the tightness of the bond between the coal seam and the surrounding rock, affecting the overall stability and bearing capacity of the coal pillar. It is an important material parameter for calculating the stress distribution in the plastic zone of the coal pillar.

[0017] The thickness of each rock layer refers to the thickness of each rock layer above the coal seam, measured in meters. It is used to identify the distribution characteristics of the immediate roof and basic roof rock layers. It is a basic geometric parameter for determining the rock layer structure, calculating the bearing capacity of the basic roof rock layer, and selecting the target layer for roof cutting.

[0018] Rock expansion coefficient This is the ratio of the volume of the crushed rock layer to the original rock volume. It is a dimensionless parameter. This coefficient reflects the degree of looseness of the rock layer after crushing. It is used to calculate the critical thickness of each basic top rock layer entering the fracture zone. It is a key parameter for determining the boundary between the caving zone and the fracture zone and selecting the target layer for cutting the top.

[0019] Furthermore, the boundary position of the plastic zone is calculated based on the geological parameters of the coal pillar. , including: calculating the vertical stress at each point in the plastic zone under coal seam compression conditions based on coal seam geological parameters : ,in, is the vertical stress at each point in the plastic zone, f is the coal body strength coefficient, and x is the distance into the coal wall. This formula can quantitatively predict the stress state at any position inside the coal pillar, providing an accurate mechanical criterion for determining whether the coal body has entered plastic deformation.

[0020] According to the vertical stress , set the boundary condition where the support pressure reaches the maximum value at the junction of the elastic-plastic zone before the roof of the goaf boundary breaks, that is, when hour, ,in, The maximum support pressure is the distance into the coal wall; in particular, when the vertical stress on the coal body reaches When the coal body begins to change from elastic state to plastic state, the establishment of this critical criterion provides a clear physical basis for determining the boundary of the plastic zone, avoiding the subjectivity and uncertainty of traditional empirical judgment.

[0021] According to the determined boundary conditions and the collected coal pillar geological parameters, the relationship coefficient between the maximum support pressure of the section coal pillar and the coal seam strength is calculated. : ,in, is the maximum stress concentration factor of the coal seam; When , it indicates that the support pressure exceeds the bearing limit of the coal body, and the coal body will inevitably enter the plastic deformation state; when When the coal body can still maintain the elastic state. According to the boundary conditions and vertical stress , calculate the boundary position of the plastic zone of the coal pillar in the section : .

[0022] In particular, when the range of the plastic zone of the coal pillar is accurately determined, the top cutting and tunnel protection technology can be used to reduce the pressure transmission of the overburden to the coal pillar and reduce the stress concentration factor. , so that the plastic zone boundary The concrete wall shrinks towards the goaf, reducing the scope of coal pillar damage. At the same time, the concrete wall designed based on accurate plastic zone analysis can effectively bear the surrounding rock pressure in the crushed area and reduce the bearing burden of the coal pillar.

[0023] Furthermore, according to the location of the plastic zone boundary The coal pillar of the section is divided into the crushed zone, the fractured zone and the original structure zone, including: starting from the edge of the goaf, the coal pillar is divided into three areas along the horizontal direction of the section coal pillar:

[0024] The crushed zone, from the edge of the goaf to the boundary of the crushed zone, is where the coal has completely broken and lost its bearing capacity. The crushed zone refers to the area from the edge of the goaf to the boundary of the crushed zone where the coal has completely broken and lost its bearing capacity due to high stress. The coal in this area has lost its original structural integrity and self-bearing capacity, and is primarily in a loose and crushed state, unable to provide effective surrounding rock support.

[0025] From the crushing zone to the boundary of the plastic zone The fracture zone is the area from the boundary of the crushing zone to the boundary of the plastic zone. Within this range, the coal body undergoes plastic deformation but still has some bearing capacity. Although the coal body in this area has cracks and local damage, the overall structure has not completely failed and can still bear a certain load, providing limited support function.

[0026] From the plastic zone boundary position The original structural zone within the coal pillar maintains its original structure and full bearing capacity. This zone extends from the plastic zone boundary x0 into the coal pillar. Within this zone, the coal maintains its original structure and full bearing capacity. This area is less affected by working face mining, and the coal's physical and mechanical properties remain largely unchanged, providing stable and reliable bearing support.

[0027] Furthermore, according to the geological parameters of the coal pillar, the basic top rock layer is determined, the bearing capacity of each basic top rock layer is calculated, the basic top rock layer with the smallest bearing capacity is used as the top cutting target layer, and the hydraulic fracturing top cutting drilling depth required to penetrate the target layer is determined, including: according to the thickness of each rock layer, the rock layer structure is identified in the order from the coal seam upwards, and the first layer above the coal seam with a thickness less than a threshold is selected. And the uniaxial compressive strength is lower than the threshold The rock layer with a thickness greater than the threshold value is taken as the direct top. And the uniaxial compressive strength is greater than the threshold The rock layers are sequentially designated as the i-th basic roof layer. The basic roof layer refers to the thickest and highest uniaxial compressive strength rock layer directly above the coal seam. It is the key bearing layer controlling the stability of the goaf roof and transmitting stress to the coal pillar. The fracture and movement of the basic roof layer directly affect the stress state and deformation characteristics of the coal pillar.

[0028] According to the identified basic top rock layers and rock layer expansion coefficient , calculate the critical thickness of the i-th basic top rock layer entering the fracture zone from the bottom up ;

[0029] According to the calculated critical thickness of each basic top rock layer Compare the thickness of the rock layer with the actual rock layer collected to determine the highest basic top rock layer entering the fracture zone. The part above the highest basic top rock layer is regarded as the fracture zone rock layer, and the part below is regarded as the caving zone rock layer. ,in, is the critical thickness of the i-th basic top rock layer from bottom to top entering the fracture zone; is the thickness of the basic top rock layer of the i-th layer from bottom to top; M is the mining height of the coal seam; is the rock expansion coefficient of the basic top rock layer; is the rock expansion coefficient of the immediate top rock layer; is the thickness of the direct roof; specifically, when the bottom surface height of a basic top rock layer exceeds the height M of the goaf, the rock layer cannot collapse into the goaf and can only bend and sink to form a fracture zone.

[0030] The top rock layer at the boundary between the fracture zone and the caving zone is used as the target layer for top cutting. The basic drilling depth is determined by the vertical distance from the coal seam roof to the target layer. The target layer for top cutting refers to the specific top rock layer that needs to be cut off by hydraulic fracturing, determined through bearing capacity analysis. The top rock layer with the lowest bearing capacity is usually selected as the target for top cutting. Cutting off this layer can effectively destroy the continuity of the overlying rock layer and reduce the stress transmitted downward.

[0031] The hydraulic fracturing topping depth is calculated based on the base drilling depth. A penetration distance is set above the base drilling depth. The penetration distance is 1 / N of the thickness of the adjacent rock formation above the topping target layer; N is a value ranging from 2 to 3. The sum of the base drilling depth and the penetration distance is used as the hydraulic fracturing topping depth, so that the hydraulic fracturing can fully penetrate the topping target layer and form a fracture network.

[0032] The hydraulic fracturing top-cutting drilling depth refers to the total length of the borehole required to penetrate the target top-cutting layer and extend the penetration distance upward from the coal seam roof. This depth ensures that hydraulic fracturing can completely cut through the target layer and form an effective fracture network above it, achieving the desired pressure relief effect.

[0033] In particular, the immediate roof stratum is thin and weak, and will collapse directly after the goaf is formed. However, the basic roof stratum is thick and strong, and has the bearing capacity to span the goaf. The thicker rock layer has greater bending stiffness and can bear greater span loads. It is the key layer for controlling the stability of the goaf and transmitting stress to the coal pillar.

[0034] On the other hand, the top cutting and tunnel protection technology changes the stress transfer mode of the overburden by disconnecting the key bearing layer. The stress originally transmitted to the coal pillar through the continuous rock layer undergoes a "stress interruption" at the top cutting position, making it impossible for the far-field stress to be effectively transmitted to the adjacent tunnel. Combined with the distribution of the plastic zone of the coal pillar determined in step S2, the degree of improvement of the stress state of the coal pillar after top cutting can be predicted, providing more accurate load boundary conditions for concrete wall reinforcement. When the key basic top rock layer is cut off, the overburden loses its continuous bearing capacity, and the stress concentration factor Significantly reduced, according to the calculation formula of the plastic zone boundary It can be seen that with As the value decreases, the plastic zone x0 shrinks accordingly. At the same time, severing the continuity of the rock formation reduces the transmission of lateral stresses and improves the stress environment in adjacent tunnels. This fundamentally solves the problem of deformation and instability in the surrounding rock of adjacent tunnels on the side of the coal pillar under deep mining conditions.

[0035] Furthermore, in S4, the range of the crushing zone is used as the influence range of the surrounding rock pressure calculation, and the static load q imposed by the surrounding rock pressure on the concrete wall is calculated. The static load q is corrected by the dynamic load coefficient to obtain the support load Q borne by the concrete wall. Correcting the static load by the dynamic load coefficient to obtain the support load Q includes: calculating the static load q borne by the concrete wall based on the range of the crushing zone and geological parameters: , where: q represents the concrete wall support load; x represents the concrete wall width; represents the outer cantilever distance of the concrete wall support; γ represents the bulk density of the separated rock blocks on the top plate; h represents the mining height; θ represents the shear angle; represents the inclination of the coal seam; H represents the height of the roof collapse; the dynamic load coefficient η is set to generate the dynamic pressure on the direct top rock layer caused by the rotation of the basic top rock layer; the dynamic load coefficient η ranges from 1 to 3, and the support load Q borne by the concrete wall is obtained based on the dynamic load coefficient η and the static load q.

[0036] Surrounding rock pressure refers to the compressive load exerted on the concrete wall by the unstable surrounding rock within the crushing zone. This pressure includes the self-weight of the crushed coal and the additional pressure transmitted to the concrete wall by the overlying rock through the crushing zone. This pressure is the primary basis for concrete wall load-bearing design and reflects the load requirements imposed on the artificial support structure by surrounding rock that has lost its self-bearing capacity.

[0037] The crushing zone range refers to the area from the edge of the goaf to the boundary of the crushing zone, where the coal body is completely crushed and loses its bearing capacity, as determined by the plastic zone analysis in step S2. This range defines the specific spatial location where the concrete wall must bear the surrounding rock pressure and is a key boundary condition for calculating the support load.

[0038] The outer overhang distance (bc) of the concrete wall support is the horizontal distance between the outer edge of the concrete wall and the edge of the roof on the goaf side, measured in meters. This parameter determines the extent of the overhanging roof that the concrete wall needs to support and directly affects the magnitude and distribution of the surrounding rock load borne by the concrete wall.

[0039] Roof detachment refers to rock fragments that have separated from the intact roof due to the loss of support from the underlying coal mass. These rock fragments, having lost their original bearing foundation, must be borne by artificial support structures such as concrete walls, contributing significantly to the surrounding rock pressure.

[0040] Roof collapse height (H) refers to the total height of rock strata that have collapsed upward from the coal seam roof within the goaf, measured in meters. This height reflects the extent of rock collapse above the goaf, affects the path of stress transfer from the overlying rock strata to the concrete wall, and influences the load magnitude. It is a key parameter for calculating surrounding rock pressure.

[0041] The rotation of the basic roof stratum refers to the rotational movement of the basic roof stratum toward the goaf, which occurs after the working face is mined, with the fracture point as the fulcrum. This rotation exerts compression and impact on the underlying direct roof stratum, and is the main mechanism for generating dynamic load effects.

[0042] The dynamic load coefficient η refers to the correction coefficient that takes into account the dynamic impact load on the concrete wall caused by mining processes such as the rotation of the basic top rock layer. It is a dimensionless parameter with a value range of 1-3.

[0043] Compared with the existing technology, the advantages of this application are:

[0044] (1) By establishing the vertical stress distribution of each point in the plastic zone during the coal seam loading process, combined with the support pressure boundary conditions at the junction of the elastic-plastic zone, the boundary position of the plastic zone is accurately calculated, and the crushing zone, fracture zone and original structure zone of the coal pillar are scientifically divided. Combined with the top cutting and tunnel protection technology of step S3, the basic top rock layer at the boundary between the fracture zone and the caving zone is cut through hydraulic fracturing to destroy the continuity and integrity of the overburden, and reduce the lateral constraint and pressure transmission of the overburden on the coal pillar. When the overburden loses its continuous bearing capacity, the vertical stress and lateral stress acting on the coal pillar are significantly reduced. According to the calculation formula of the plastic zone boundary, as the stress concentration coefficient increases, the vertical stress and lateral stress acting on the coal pillar decrease significantly. The decrease of plastic zone The width of coal pillar damage is effectively controlled.

[0045] (2) The optimal top cutting target layer and drilling depth are determined through rock layer structure analysis. The rock separation method and dynamic load coefficient correction are combined to accurately calculate the concrete wall support load requirement, and the top cutting and tunnel protection technology is organically combined with the concrete wall reinforcement technology. Compared with the existing technology of applying the two technologies independently, this application can reduce the width of the coal pillar damage on the goaf side, enhance the bearing capacity of the coal pillar, and weaken the lateral mining stress transmission through collaborative design.

[0046] (3) The top-cutting tunnel protection and the concrete wall reinforcement form a mechanical coupling mechanism, blocking the stress transmission path from two dimensions. First, the top-cutting tunnel protection breaks the continuity of the rock layer in the target layer, causing the horizontal stress originally transmitted along the rock layer to concentrate and release at the top-cut position, thereby blocking the direct transmission path of the far-field stress to the adjacent tunnel. Second, the setting of the concrete wall changes the near-field stress distribution pattern. According to the wall thickness and layout position determined in step S5, the concrete wall forms a new stress transmission medium between the coal pillar and the goaf. Since the elastic modulus of concrete is much larger than that of the broken coal body, the stress is preferentially transmitted through the concrete wall, reducing the stress component transmitted through the coal pillar to the adjacent tunnel. The synergistic effect of the two technologies enables the lateral mining stress to be effectively dispersed and absorbed during the transmission process. BRIEF DESCRIPTION OF THE DRAWINGS

[0047] The present application will be further described in the form of exemplary embodiments, which will be described in detail with reference to the accompanying drawings. These embodiments are not limiting, and in these embodiments, the same numbers represent the same structures, wherein: Figure 1 This is an exemplary flow chart of a method for reinforcing a concrete wall on the side of a coal pillar goaf in a mining area according to some embodiments of the present application; Figure 2 is a schematic diagram of the roof rock structure according to some embodiments of the present application; Figure 3 is a schematic diagram of a coal seam structure along a void according to some embodiments of the present application; Figure 4 is a schematic diagram of the top plate structure and vertical stress distribution according to some embodiments of the present application; Figure 5 This is a process flow chart for constructing a concrete wall using the spraying process shown in some embodiments of the present application. DETAILED DESCRIPTION

[0048] The method and system provided in the embodiments of the present application are described in detail below with reference to the accompanying drawings.

[0049] The mine covers an area of ​​10.8186 square kilometers, with a production capacity of 1.2 million tons per year. Mining depths range from 900 meters to 300 meters above seam level. The mine utilizes central parallel ventilation. The mine is divided into two levels: Level 1, which mines the No. 3 coal seam, and Level 2, which mines the No. 15 coal seam. The mine is currently mining the No. 3 coal seam on Level 1, at an elevation of +663 meters. The mining area is divided into three panels: Panels 1, 2, and 3, from north to south. The primary mining area is the center of Panel 1.

[0050] According to the geological description of the Yipan area and the mining status information provided by the mine, the mine is currently carrying out mining activities in the Yipan area (east). The Yipan area (east) is mined on two wings, with 3101, 3103, 3105, 3107, and 3109 working faces arranged on the west wing, and 3102, 3104, 3106, 3108, and 3110 working faces arranged on the east wing.

[0051] The coal seam in the first panel area of ​​Xiaoxi Coal Mining is approximately 600-700 meters deep. This is a typical deep-shaft mining method, employing full-height mining (maximum mining height), single-wing sequential mining, and dual-lane tunneling. The coal pillar width is 25 meters. Mining operations within this working face significantly disrupt adjacent lanes. For example, during mining operations at the 3108 fully mechanized mining face, the gas control lane in the east wing suffered significant deformation.

[0052] The direct roof of the tunnel is mostly fine-grained sandstone and sandy mudstone, with some parts being silty mudstone, mudstone or muddy siltstone. The rock mass has low strength and is easily weathered and broken. The tunnel design has a large cross-section, making support difficult.

[0053] According to the "Test Report on the Physical and Mechanical Parameters of the Roof Rock Stratum of East No. 3 Coal Seam in the First Pan Area of ​​Xiaoxi Coal Industry Co., Ltd., Yangcheng Yangtai Group, Shanxi Province", the uniaxial compressive strength of coal is 10.4~14.5Mpa, and the strength of the coal body is low. After the working face is mined, the coal pillars protecting the tunnel will undergo plastic deformation due to the influence of support pressure.

[0054] The first plate area has many structures, poor integrity, weakened self-bearing capacity, and the surrounding rock of the tunnel becomes loose and broken, causing the anchor rods (cables) to lose their stable support basis and unable to play an active support role. Under the combined action of rock expansion pressure and deformation pressure, they flow and squeeze into the tunnel, forming net bags and slag falling phenomena. As the degree and depth of surrounding rock fragmentation continue to increase, the anchor rods (cables) gradually lose their support capacity, and the deformation and instability of the tunnel intensify.

[0055] In view of the above reasons, although a variety of comprehensive measures have been taken to control the gas control lanes in the east wing of the first plate area, serious deformation still occurs while the cost increases (large movement of the roof and floor plates, convergence of the two sides, support failure, etc.). The traditional coal pillar support and lane reinforcement technology has limited effect. Therefore, under the original comprehensive control, it is planned to carry out the research on the mechanical coupling mechanism and coordinated bearing technology of the concrete wall reinforcement on the side of the coal pillar goaf in the section to prevent the subsequent mining disturbance of the working face from causing serious deformation of the adjacent lanes. Figure 1 As shown, the geological parameters of the coal pillar section are collected. The coal pillar section is the coal body between the goaf formed by the working face mining and the area to be mined. According to the geological parameters of the coal pillar, the boundary position of the plastic zone is calculated. , according to the position of the plastic zone boundary The coal pillar of the section is divided into a crushed zone, a fractured zone and an original structure zone; according to the geological parameters of the coal pillar, the basic top rock layer is determined, the bearing capacity of each basic top rock layer is calculated, the basic top rock layer with the smallest bearing capacity is used as the top cutting target layer, and the hydraulic fracturing top cutting drilling depth required to penetrate the target layer is determined; the range of the crushed zone is used as the influence range of the surrounding rock pressure calculation, the static load q imposed by the surrounding rock pressure on the concrete wall is calculated, and the static load q is corrected by the dynamic load coefficient. The static load is corrected by the dynamic load coefficient to obtain the support load Q; according to the support load Q, the concrete strength grade of the concrete wall, the thickness of the concrete wall in the direction perpendicular to the tunnel axis in the tunnel section, and the horizontal distance between the top of the concrete wall and the edge of the roof of the goaf are determined.

[0056] Specifically, during coal mining, the roof rock layer is disturbed and undergoes different stress field evolution and rebalancing processes due to changes in the supporting structure between the rock layers. The roof rock layers at different locations in the disturbed area will be damaged to varying degrees and exhibit different structural characteristics, such as Figure 2 As shown, three horizontal zones (A-coal seam support zone, B-separation fracture zone, C-broken collapse zone) and three vertical zones (Ⅰ-caving zone, Ⅱ-fissure zone, Ⅲ-bending subsidence zone) are formed.

[0057] After the coal body along the edge of the goaf undergoes the evolution of support pressure, it often presents the characteristics of crushing-fracture-original structure from the edge to the inside (such as Figure 3The edge crushing and internal cracking are caused by the strength failure during the redistribution of the supporting pressure formed by the working face mining. The size of the crushed and cracked areas is affected by factors such as coal body properties, rock layer structure, mining thickness, goaf size, and mining depth.

[0058] The yield line analysis method is used to infer the roof yield failure line in the longwall mining face. In the early stage of mining, the roof will form an "O-X" fracture. In the subsequent mining process of the working face, internal stress fields and external stress fields will be formed in front of and on the side of the working face. The lateral roof structure and vertical stress distribution of the working face are as follows: Figure 4 shown.

[0059] Coal seams produce plastic zones during loading, and the internal stress field is usually distributed in this zone. Considering the compression of the coal seam, the vertical stress expression of each point in the plastic zone is:

[0060] (1)

[0061] Where: is the vertical stress at each point in the plastic zone; is the unidirectional compressive strength of the coal seam; C is the cohesion between the coal seam and the roof and floor; f is the solidity coefficient of the coal body; h is the mining thickness of the coal seam; is the coal seam compression angle; x is the distance into the coal wall; The distance into the coal wall where the maximum support pressure is reached.

[0062] Before the roof of the goaf boundary breaks, the support pressure at the junction of the elastic-plastic zone reaches its maximum value, that is, when hour,

[0063] (2)

[0064] in: is the maximum stress concentration coefficient of the coal seam. Substituting this condition into formula (1) can obtain the range of the plastic zone: The expression is:

[0065] (3)

[0066] in: is the relationship coefficient between the maximum support pressure and the coal seam strength condition, and its expression is:

[0067] (4)

[0068] It can be seen that the burial depth H and the maximum stress concentration factor The larger the value is, the larger the plastic zone is. The smaller, The bigger.

[0069] Based on the above analysis, it can be concluded that under a certain mining depth and given coal seam conditions, the range of the plastic zone is proportional to the coal seam mining thickness h, and the specific boundary of the internal stress field is determined by the rock beam strength and the clamping force required for fracture. The greater the rock beam mass and the higher the strength, the greater the clamping force required for fracture and the larger the range of the plastic zone. Before the lateral roof fractures, a lateral peak stress will be formed at the junction of the elastic and plastic zones of the coal pillar. At this stage, the tunnel will be affected by the lateral stress, which will cause large deformation and shrinkage of the tunnel cross-section.

[0070] Therefore, in order to alleviate the deformation of goaf-side tunnels, research on the top cutting and tunnel protection technology and the mechanical coupling and collaborative bearing technology of the concrete wall reinforcement on the goaf side of the coal pillar should be carried out in the mining tunnel. The purpose is to reduce the damage width of the coal pillar on the goaf side, enhance the bearing capacity of the coal pillar, weaken the transfer of lateral mining stress to the adjacent tunnel, and improve the stress environment of the adjacent tunnel.

[0071] Calculation of hydraulic fracturing top cutting and tunnel protection, the required top cutting height of the working face:

[0072] According to the distribution characteristics of No. 3 coal stratum in Xiaoxi Coal Mine:

[0073]

[0074] The basic top rock layer entering the fracture zone is calculated according to the following formula:

[0075] , is the thickness of the i-th foundation rock layer from bottom to top; is the thickness of the basic top layer of the i-th layer from bottom to top; M- is the mining height of the coal seam; is the rock expansion coefficient of the basic roof and its additional rock layers; is the rock expansion coefficient of the immediate roof and its additional rock layers; h is the thickness of the immediate roof;

[0076] Substituting the working face data into the above formula reveals the third basic roof ridge, meaning that the strata above the 16.84m sandy mudstone are the fracture zone, and the strata below are the caving zone. Therefore, the vertical depth of the fracturing drill hole should reach above the third basic roof ridge. To ensure effective fracturing, the fracturing position should be as deep as possible, penetrating half the thickness of the fourth fine-grained sandstone layer.

[0077] In summary: the required top-breaking height for cutting along the air is 40m.

[0078] Based on the geological and mining conditions at the Xiaoxi Coal Mine working face, the "separated rock block method" was chosen to calculate the surrounding rock pressure on the concrete wall. The separated rock block method, a theoretical method for calculating support loads based on the mechanism of surrounding rock instability, is used to determine the magnitude of the surrounding rock pressure that the concrete wall must bear. The core principle of this method is to treat the rock masses within a certain influence range above the concrete wall, separated from the intact surrounding rock due to the collapse of the underlying coal mass and the loss of support, as independent load units for analysis and calculation. Specifically, the separated rock block method posits that when the coal pillar collapses and loses its bearing capacity, the roof strata above it separate and fracture under the influence of gravity and overburden pressure, forming relatively independent rock blocks. The weight of these separated rock blocks, along with the overburden pressure transmitted through them, constitutes the primary static load borne by the concrete wall. This method establishes a geometric model of the separated rock blocks and, incorporating parameters such as their bulk density, thickness, and distribution range, applies the principle of mechanical equilibrium to calculate the surrounding rock pressure acting on the concrete wall.

[0079] The theoretical basis of this method is that the weight of the separated rock blocks within a certain range above the concrete wall constitutes the static load of the support. The formula for calculating the rock pressure of the concrete wall is as follows:

[0080]

[0081] Where: —Concrete wall support load;

[0082] —Concrete wall width, take 3m;

[0083] —The outer cantilever distance of the concrete wall support is 1.5m;

[0084] —Bulk density of separated rock blocks on the roof, take 27kN / m3;

[0085] - Mining height: 6.2m;

[0086] — Shear angle, selected as 26° based on experience;

[0087] —coal seam inclination, 2°;

[0088] —The height of roof collapse. If the rock layer expansion coefficient is 1.2, the height of roof collapse is 31m.

[0089] It is calculated that when the support width is 3m, the pressure on the tunnel side support is

[0090] q=6.5MPa

[0091] The load on the concrete wall comes from the weight of the separated rock blocks on the immediate roof and the dynamic pressure caused by the rotation of the old roof on the immediate roof. The dynamic pressure caused by the rotation of the old roof on the immediate roof has always been difficult to determine. The traditional view is to measure the so-called dynamic load coefficient η by actual measurement. According to the empirical data of fully mechanized mining, η is generally 1 to 2. When designing the concrete wall support, it is set to 2. The maximum load on the concrete wall is:

[0092] 2 =2×6,5=13MPa

[0093] In summary, the concrete strength grade used for the concrete wall support on the goaf side of the coal pillar in the Xiaoxi Coal Mine section is C30, which can meet the strength requirements.

[0094] Spraying technology is an innovative method for constructing concrete walls. Sand, cement and JCT-1 spray concrete additive are mixed in a specific proportion in a wet (or dry) environment and then placed in a spraying device. The mixture is transported to the nozzle through a pipe using compressed air. After being combined with water, it is sprayed onto the construction surface at high speed to form a coating or wall. The specific process is as follows: Figure 5 The characteristics of using the spraying process to construct concrete walls or layers are as follows: Only temporary barriers need to be set up during construction, and a thick layer can be achieved in a single spraying operation. The wall quickly takes shape, eliminating the need for support structures and subsequent maintenance. The sprayed surface has strong adhesion, effectively suppressing dust and trapping moisture, with a spray rebound rate of less than 5%. Initial hardening is rapid, with no decrease in maximum hardness and an ultimate strength exceeding C30. The grafting performance is excellent, with the sprayed layer and rock stratum sharing the load. The wall is characterized by high density, high strength, minimal expansion, a compact texture, and airtightness. The construction machinery used is compact and quick and easy to operate, reducing costs and manpower requirements, and achieving relatively fast construction speeds.

[0095] Using heap shotcrete technology to construct retaining walls is not only fast and easy to construct, but also ensures concrete strength, offers significant cost advantages, and fully complies with safety standards for underground coal mining operations. The speed of wall construction with heap shotcrete technology depends primarily on the output power of the shotcrete machinery. If the equipment performs well, the process can achieve an efficiency of 7-8 cubic meters per hour. This plays a key role in increasing wall construction speed, reducing labor intensity, and shortening construction time, making it crucial for promoting safe and efficient coal mine operations.

[0096] The invention of the present application and its implementation methods are described schematically above. This description is not restrictive. Without departing from the spirit or basic features of the present application, the present application can be implemented in other specific forms. What is shown in the accompanying drawings is only one of the implementation methods of the invention of the present application, and the actual structure is not limited to this. Therefore, if a person of ordinary skill in the art is inspired by it, without departing from the purpose of the invention, a structural method and embodiment similar to the technical solution are designed without creativity, which should all fall within the scope of protection of the present application. In addition, the word "including" does not exclude other elements or steps, and the word "one" before an element does not exclude the inclusion of "multiple" elements. Words such as first and second are used to indicate names and do not indicate any specific order.

Claims

1. A method for reinforcing the concrete wall of the coal pillar goaf side of a mining section, characterized in that: include: Collecting geological parameters of the section coal pillar, where the section coal pillar is the coal body between the goaf formed by mining at the working face and the area to be mined; Calculate the boundary position of the plastic zone based on the geological parameters of the coal pillar , according to the position of the plastic zone boundary The coal pillars in the section are divided into crushed zone, fractured zone and original structure zone; According to the geological parameters of the coal pillar, the basic top rock layer is determined, the bearing capacity of each basic top rock layer is calculated, the basic top rock layer with the smallest bearing capacity is used as the top cutting target layer, and the hydraulic fracturing top cutting drilling depth required to penetrate the target layer is determined; The crushing area is taken as the influence range of the surrounding rock pressure calculation, and the static load q exerted by the surrounding rock pressure on the concrete wall is calculated. The static load q is corrected by the dynamic load coefficient, and the support load Q is obtained by correcting the static load by the dynamic load coefficient. According to the support load Q, determine the concrete strength grade of the concrete wall, the thickness of the concrete wall in the direction perpendicular to the tunnel axis in the tunnel section, and the horizontal distance between the top of the concrete wall and the edge of the side roof of the goaf.

2. The method for reinforcing the concrete wall of the coal pillar goaf side of the mining section according to claim 1 is characterized in that: The coal pillar geological parameters include: uniaxial compressive strength of the coal seam , the average bulk density γ of the overlying rock layer, the buried depth H, the coal seam mining thickness h, the coal seam compression angle α, the cohesion C between the coal seam and the roof and floor, as well as the thickness of each rock layer and the rock layer expansion coefficient .

3. The method for reinforcing the concrete wall of the coal pillar goaf side of the mining section according to claim 2 is characterized in that: Calculate the location of the plastic zone boundary ,include: According to the geological parameters of the coal seam, calculate the vertical stress at each point in the plastic zone under the compression condition of the coal seam : ;in, is the vertical stress at each point in the plastic zone, f is the coal body solidity coefficient, and x is the distance into the coal wall; is the unidirectional compressive strength of the coal seam; According to the vertical stress , set the boundary condition where the support pressure reaches the maximum value at the junction of the elastic-plastic zone before the roof of the goaf boundary breaks, that is, when hour, ,in, The maximum support pressure depth into the coal wall; According to the determined boundary conditions and the collected coal pillar geological parameters, the relationship coefficient between the maximum support pressure of the section coal pillar and the coal seam strength is calculated. : ;in, is the maximum stress concentration factor of the coal seam; According to the boundary conditions and vertical stress , calculate the boundary position of the plastic zone of the coal pillar in the section : .

4. The method for reinforcing the concrete wall of the coal pillar goaf side of the mining section according to claim 3 is characterized in that: According to the position of the plastic zone boundary The coal pillars in the section are divided into crushed zone, fractured zone and original structure zone, including: Starting from the edge of the goaf, the coal pillar is divided into three areas along the horizontal direction of the coal pillar: The crushing zone from the edge of the goaf to the boundary of the crushing zone where the coal body is completely crushed and loses its bearing capacity; From the crushing zone to the boundary of the plastic zone The fracture zone within the range where the coal body undergoes plastic deformation but still has some bearing capacity; From the plastic zone boundary position The original structural zone within the coal pillar where the coal body maintains its original structure and complete bearing capacity.

5. The method for reinforcing the concrete wall of the coal pillar goaf side of the mining section according to claim 3 is characterized in that: Determine the hydraulic fracturing cut-off drilling depth required to penetrate the target formation, including: According to the thickness of each rock layer, the rock layer structure is identified in the order from the coal seam upwards, and the first layer above the coal seam with a thickness less than the threshold is selected. And the uniaxial compressive strength is lower than the threshold The rock layer with a thickness greater than the threshold value is taken as the direct top. And the uniaxial compressive strength is greater than the threshold The rock layer is sequentially used as the i-th basic top rock layer; According to the identified basic top rock layers and rock layer expansion coefficient , calculate the critical thickness of the i-th basic top rock layer entering the fracture zone from the bottom up ; According to the calculated critical thickness of each basic top rock layer Compare the thickness of the rock layer with the actual rock layer collected to determine the highest basic top rock layer entering the fracture zone. The part above the highest basic top rock layer is regarded as the fracture zone rock layer, and the part below is regarded as the caving zone rock layer. The basic top rock layer at the boundary between the fracture zone and the caving zone is taken as the target layer for cutting the top, and the basic drilling depth is determined according to the vertical distance from the coal seam roof to the target layer for cutting the top; The hydraulic fracturing top cutting drilling depth is calculated based on the basic drilling depth.

6. The method for reinforcing the concrete wall of the coal pillar goaf side of the mining section according to claim 5, characterized in that: Calculate the critical thickness of the i-th basic top rock layer from bottom to top to enter the fracture zone , using the following formula: ;in, is the critical thickness of the i-th basic top rock layer from bottom to top entering the fracture zone; is the thickness of the basic top rock layer of the i-th layer from bottom to top; M is the mining height of the coal seam; is the rock expansion coefficient of the basic top rock layer; is the rock expansion coefficient of the immediate top rock layer; is the direct top thickness.

7. The method for reinforcing the concrete wall of the coal pillar goaf side of the mining section according to claim 5, characterized in that: The hydraulic fracturing top cutting drilling depth is calculated based on the basic drilling depth, including: A penetration distance is set above the basic drilling depth, where the penetration distance is 1 / N of the thickness of the adjacent rock layer above the top cutting target layer; The sum of the basic drilling depth and the penetration distance is taken as the hydraulic fracturing top cutting drilling depth, so that the hydraulic fracturing can completely penetrate the top cutting target layer and form a fracture network.

8. The method for reinforcing the concrete wall of the coal pillar goaf side of the mining section according to claim 7, characterized in that: The value range of N is 2 or 3.

9. The method for reinforcing the concrete wall of the coal pillar goaf side of the mining section according to claim 7, characterized in that: Obtain the support load Q borne by the concrete wall; The static load is modified by the dynamic load coefficient to obtain the support load Q, which includes: According to the scope of the crushing area and geological parameters, the static load q borne by the concrete wall is calculated: ； Where: q represents the concrete wall support load; x represents the concrete wall width; represents the outer cantilever distance of the concrete wall support; γ represents the bulk density of the separated rock blocks on the roof; h represents the mining height; θ represents the shear angle; Indicates the inclination of the coal seam; H indicates the height of the roof collapse; Set the dynamic load coefficient η of the dynamic pressure caused by the rotation of the basic top rock layer on the direct top rock layer; According to the dynamic load coefficient η and the static load q, the support load Q borne by the concrete wall is obtained.

10. The method for reinforcing the concrete wall of the coal pillar goaf side of the mining section according to claim 9, characterized in that: The dynamic load coefficient η ranges from 1 to 3.

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