A method for reinforcing a concrete wall on the goaf side of a coal pillar in a mining section

Abstract

The application discloses a kind of exploitation section coal pillar goaf side concrete wall reinforcing method, it is related to the field of coal mining, comprising: collecting section coal pillar geological parameters;According to the coal pillar geological parameters, the position of plastic zone boundary is calculated, according to section coal pillar is divided into broken zone, broken zone and original structure zone;According to the coal pillar geological parameters, determine the basic roof rock, calculate the bearing capacity of each basic roof rock, with the smallest basic roof rock as the target horizon of cutting roof, determine the water pressure fracturing cutting roof drilling depth required for penetrating target horizon;The range of broken zone is used as the influence range of surrounding rock pressure calculation, the static load q exerted by surrounding rock pressure on concrete wall is calculated, the dynamic load coefficient is used to correct static load q, and the supporting load Q is obtained;According to supporting load Q, determine the reinforcing parameter.The application effectively controls the deformation and instability of surrounding rock caused by mining disturbance of working face in the section coal pillar goaf side adjacent roadway under the condition of deep well mining.

Description

Technical Field

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

[0002] With the continuous deepening of coal resource mining in my country, shallow coal resources are becoming increasingly depleted, and coal mining is gradually shifting towards deeper areas. Deep 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 mining conditions, problems such as high ground stress, complex rock mass properties, and strong disturbance effects are becoming increasingly prominent, posing severe challenges to safe and efficient coal mining. Especially in deep mining areas using high-extraction, longwall mining techniques, the disturbance effect of working face mining on adjacent roadways is more significant, and controlling the deformation of the surrounding rock in roadways has become a key technical problem restricting safe production in deep coal mines.

[0003] In deep-well mining, the section coal pillar, as the coal body between the goaf formed by the working face and the area to be mined, plays a crucial role in protecting adjacent roadways. However, due to the high ground stress and relatively low coal strength in deep wells, the section coal pillar is subjected to intense supporting pressure after the working face is mined, making it prone to plastic deformation or even failure, resulting in a significant decrease in its load-bearing capacity. Simultaneously, the roof strata in deep-well mining are often of low strength and easily weathered and fractured, and the roadway design cross-section is large, resulting in insufficient self-bearing capacity of the surrounding rock. Under geological conditions of well-developed structures and poor rock mass integrity, traditional anchor bolt support is insufficient to establish an effective load-bearing structure in loose and fractured surrounding rock, leading to phenomena such as flow extrusion and slag shedding in the roadway surrounding rock, exacerbating the problem of surrounding rock deformation and instability.

[0004] Existing coal pillar support and roadway reinforcement technologies mainly rely on passive measures such as increasing support density and strength, lacking a scientific understanding of the distribution of the plastic zone in the coal pillar and accurate calculation of the surrounding rock load. Support designs are often determined based on experience, lacking theoretical basis. Furthermore, roof cutting and roadway protection technologies and coal pillar reinforcement technologies are usually applied independently, lacking synergistic optimization design and failing to form an effective mechanical coupling and synergistic load-bearing system. Therefore, there is an urgent need to establish a scientific method for reinforcing the concrete wall on the goaf side of the coal pillar in deep mining sections. Through precise analysis of the coal pillar plastic zone, reasonable roof cutting and roadway protection design, and scientific determination of concrete wall support parameters, this method can achieve mechanical coupling and synergistic load-bearing between roof cutting and roadway protection and concrete wall reinforcement, effectively controlling the deformation and instability of the surrounding rock in adjacent roadways under deep mining conditions.

[0005] For example, related technology CN118351952A discloses a method for calculating the stability of a section coal pillar, including the following steps: S1, calculation of the bending elastic energy of the basic roof on the free side: establishing a mechanical model of the basic roof on the free side, and calculating the bending elastic energy UW of the basic roof on the free side based on the mechanical model of the basic roof on the free side; S2, calculation of the yield elastic energy of the section coal pillar: establishing a lateral stress model of the section coal pillar, obtaining the calculation formula for the lateral stress distribution of the section coal pillar, and obtaining the distribution expression of the yield elastic energy Es of the elastic zone of the section coal pillar along the width direction of the coal pillar based on the calculation formula for the lateral stress distribution of the section coal pillar; S3, using UW > Es as the criterion for the instability of the coal body in the elastic zone of the section coal pillar, analyzing and judging the stability of the section coal pillar. However, this application uses the method of overall elastic energy comparison, treating the coal pillar as a homogeneous body for analysis, ignoring the complexity of the stress distribution inside the coal pillar. Summary of the Invention

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

[0007] This application provides a method for reinforcing the concrete wall on the goaf side of a mining section coal pillar, comprising: collecting geological parameters of the coal pillar, wherein the coal pillar is the coal body between the goaf formed by the working face and the area to be mined; and calculating the location of the plastic zone boundary based on the geological parameters of the coal pillar. According to the location of the plastic zone boundary The coal pillar section is divided into a fractured zone, a cracked zone, and a zone with original structure; among them, the boundary of the plastic zone is located... The stress level of a coal pillar refers to the horizontal distance from the edge of the goaf to the critical point where the coal body transitions from a plastic deformation state to an elastic state. This location marks the boundary of the coal pillar's stress state. The support pressure at the location reaches its maximum value, and the coal body beyond that location still maintains its elastic state and complete load-bearing capacity.

[0008] Based on the geological parameters of the coal pillar, the basic roof strata are determined, and the bearing capacity of each basic roof strata is calculated. The basic roof stratum with the smallest bearing capacity is selected as the target stratum for roof cutting, and the hydraulic fracturing borehole depth required to penetrate the target stratum is determined. The fractured zone is used as the influence range for calculating the surrounding rock pressure, and the static load q exerted by the surrounding rock pressure on the concrete wall is calculated. The static load q is corrected using a dynamic load factor to obtain the support load Q borne by the concrete wall. The static load is corrected using a dynamic load factor to obtain the support load Q. Based on the support load Q, the concrete strength grade of the concrete wall, the thickness of the concrete wall in the direction perpendicular to the roadway axis within the roadway cross-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 elastoplastic mechanics theory and the Mohr-Coulomb failure criterion, coal pillars undergo elastoplastic deformation under support pressure after the working face is mined. Establishing a vertical stress distribution model of the plastic zone and determining its extent by combining boundary conditions essentially provides a quantitative description of the stress-strain relationship of the coal pillar. The location of the plastic zone boundary... The calculations revealed the critical position of the coal pillar transitioning from an elastic to a plastic state, and the three-zone division based on this scientifically reflects the differences in the 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 unit weight γ of the overlying strata, burial depth H, coal seam mining thickness h, coal seam compression angle α, cohesion C between the coal seam and the roof and floor, as well as the thickness of each stratum and the coefficient of stratum fragmentation. .

[0011] Among them, the uniaxial compressive strength of coal seams This refers to the maximum compressive stress a coal sample can withstand before failure under uniaxial compression conditions, measured in MPa. This parameter reflects the inherent strength characteristics of the coal body and is a key mechanical indicator for calculating the boundary location and bearing capacity of the plastic zone of a coal column, directly affecting the deformation and failure mode of the coal column under support pressure.

[0012] The average unit weight γ of the overlying strata refers to the average weight per unit volume of all strata above the coal seam, expressed in kN / m³. This parameter is used to calculate the vertical stress exerted by the overlying strata on the coal pillar. It is a fundamental parameter for determining the loading state and plastic zone distribution of the coal pillar, directly affecting the magnitude and distribution of the supporting pressure.

[0013] Burial depth H refers to the vertical distance from the surface to the roof of the coal seam, measured in meters (m). Burial depth determines the magnitude of the self-weight stress of the overlying strata that the coal seam bears, and is an important boundary condition for calculating the initial stress state of the coal pillar and the stress redistribution after the working face is mined.

[0014] The coal seam mining thickness h refers to the actual thickness of the coal seam mined at the working face, measured in meters (m). This parameter affects the geometry of the goaf and the collapse characteristics of the overlying strata, and is an important geometric parameter for calculating the range of the plastic zone of the coal pillar and the support load of the concrete wall.

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

[0016] The cohesion C between the coal seam and the roof and floor 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, and is an important material parameter for calculating the stress distribution in the plastic zone of the coal pillar.

[0017] The thickness of each stratum refers to the thickness of each stratum above the coal seam, in meters (m). It is used to identify the distribution characteristics of the immediate roof and the main roof strata, and is a fundamental geometric parameter for determining the stratum structure, calculating the bearing capacity of the main roof strata, and selecting the target stratum for roof cutting.

[0018] Rock stratum swelling coefficient This refers to the ratio of the volume of the fractured rock strata to the volume of the original rock, and is a dimensionless parameter. This coefficient reflects the degree of looseness after the rock strata are fractured and is used to calculate the critical thickness of each basic top rock stratum entering the fracture zone. It is a key parameter for determining the boundary between the caving zone and the fracture zone and for selecting the target stratum for cutting the top.

[0019] Furthermore, based on the geological parameters of the coal pillar, the location of the plastic zone boundary is calculated. This includes: calculating the vertical stress at various points in the plastic zone of the coal seam under compression conditions based on the geological parameters of the coal seam. : ,in, Let f be the vertical stress at each point in the plastic zone, f be the coal body firmness coefficient, and x be 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 vertical stress When the boundary condition is set at the junction of the elastic-plastic zone and the boundary of the goaf, where the support pressure reaches its maximum value, i.e., when... hour, ,in, The distance into the coal face where the maximum supporting pressure is measured; specifically, when the vertical stress on the coal body reaches... At this point, the coal begins to transition from an elastic to a 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 judgments.

[0021] Based on the determined boundary conditions and the collected geological parameters of the coal pillar, the relationship coefficient between the maximum bearing pressure of the coal pillar and the coal seam strength is calculated. : ,in, The maximum stress concentration factor of the coal seam; where, when When the bearing pressure exceeds the bearing capacity limit of the coal body, the coal body will inevitably enter a state of plastic deformation; when At this time, the coal body can still maintain an elastic state. This is based on boundary conditions and vertical stress. Calculate the location of the plastic zone boundary of the coal pillar section. : .

[0022] In particular, once the plastic zone of the coal pillar is precisely determined, the pressure transmission from the overlying strata to the coal pillar can be reduced through roof cutting and roadway protection techniques, thereby lowering the stress concentration factor. Thus, the boundary of the plastic region The structure contracts towards the goaf, reducing the extent of coal pillar failure. Simultaneously, the concrete wall, designed based on precise plastic zone analysis, effectively bears the surrounding rock pressure in the fractured zone, alleviating the load on the coal pillar.

[0023] Furthermore, based on the location of the plastic zone boundary The coal pillar section is divided into three zones: a fractured zone, a fractured zone, and a zone with original structure. This includes dividing the coal pillar into three areas along the horizontal direction of the goaf edge:

[0024] The fractured zone is the area from the edge of the goaf to the boundary of the fractured zone where the coal body is completely broken and has lost its load-bearing capacity. The fractured zone refers to the area from the edge of the goaf to the boundary of the fractured zone where the coal body has completely broken and lost its load-bearing capacity under high stress. The coal body in this area has lost its original structural integrity and self-supporting capacity, mainly exhibiting a loose and broken state, and cannot provide effective surrounding rock support.

[0025] From the boundary between the fractured zone and the plastic zone The fracture zone refers to the area within which the coal body undergoes plastic deformation but still retains partial load-bearing capacity; the fracture zone is defined as the area from the boundary of the fractured zone to the boundary of the plastic zone. Within this range, the coal body undergoes plastic deformation but still retains some load-bearing capacity. Although the coal body in this area exhibits cracks and localized damage, its overall structure has not completely failed and can still bear a certain load, thus possessing limited support capabilities.

[0026] From the boundary of the plastic zone The original structural zone refers to the area extending from the boundary of the plastic zone (x0) into the coal pillar, where the coal maintains its original structure and complete load-bearing capacity. This zone is minimally affected by the face mining, and the physical and mechanical properties of the coal remain largely unchanged, providing stable and reliable load-bearing support.

[0027] Furthermore, based on the geological parameters of the coal pillar, the basic roof strata are determined, and the bearing capacity of each basic roof strata is calculated. The basic roof stratum with the lowest bearing capacity is selected as the target stratum for cutting the roof. The required hydraulic fracturing borehole depth to penetrate the target stratum is determined, including: identifying the stratum structure according to the thickness of each stratum, following the order from the coal seam upwards, and selecting the first stratum above the coal seam with a thickness less than a threshold. And the uniaxial compressive strength is below the threshold. The rock strata are taken as the immediate top, and the thickness above the immediate top is greater than the threshold. And the uniaxial compressive strength is greater than the threshold. The rock strata are sequentially designated as the i-th basic roof stratum. The basic roof stratum refers to the thickest stratum above the immediate roof of the coal seam, possessing high uniaxial compressive strength. 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 stratum directly affect the stress state and deformation characteristics of the coal pillar.

[0028] Based on the identified basic top rock layers and their strata fragmentation coefficients Calculate the critical thickness of the i-th basic top rock layer entering the fracture zone from bottom to top. ;

[0029] Based on the calculated critical thickness of each basic top rock layer By comparing the actual rock strata thickness with the actual samples, the highest basic top rock stratum entering the fracture zone is determined. The part above the highest basic top rock stratum is taken as the fracture zone rock stratum, and the part below it is taken as the caving zone rock stratum. ,in, The critical thickness for the i-th basic top rock layer to enter the fracture zone from bottom to top; M represents the thickness of the i-th basic top rock layer from bottom to top; M represents the coal seam mining height. The coefficient of rock fragmentation of the basic top stratum; The coefficient of rock fragmentation of the immediate overlying stratum; The thickness is the direct roof thickness; specifically, when the bottom height of a certain basic roof 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 basic overlying stratum at the boundary between the fracture zone and the caving zone is designated as the target stratum for hydraulic fracturing. The depth of the foundation borehole is determined based on the vertical distance from the coal seam roof to the target stratum. The target stratum refers to the specific basic overlying stratum that requires hydraulic fracturing, as determined through bearing capacity analysis. Typically, the basic overlying stratum with the lowest bearing capacity is selected as the target. Cutting this stratum effectively disrupts the continuity of the overlying strata, reducing downward stress transmission.

[0031] The hydraulic fracturing top-cutting borehole depth is calculated based on the basic borehole depth. A penetration distance is set above the basic borehole depth, which is 1 / N of the thickness of the adjacent rock layer above the target layer; N ranges from 2 to 3. The sum of the basic borehole depth and the penetration distance is taken as the hydraulic fracturing top-cutting borehole depth, so that the hydraulic fracturing can completely penetrate the target layer and form a fracture network.

[0032] The hydraulic fracturing top-cutting borehole depth refers to the total borehole length required to drill upwards from the coal seam roof, penetrating the target stratum and extending a certain penetration distance. This depth ensures that hydraulic fracturing can completely cut off the target stratum and form an effective fracture network above it, achieving the desired pressure relief effect.

[0033] Specifically, on the one hand, the immediate roof strata are thin and low in strength, and will collapse directly after the goaf is formed, while the main roof strata are thick and high in strength, and have the bearing capacity to span the goaf. Thicker strata have greater bending stiffness and can bear larger span loads, and are key strata for controlling the stability of the goaf and transmitting stress to the coal pillars.

[0034] On the other hand, the roof-cutting and roadway-protection technology alters the stress transmission pattern of the overlying strata by disconnecting the key bearing strata. The stress originally transmitted to the coal pillar through continuous strata is interrupted at the roof-cutting location, preventing effective transmission of far-field stress to adjacent roadways. Combined with the distribution of the coal pillar's plastic zone determined in step S2, the degree of improvement in the coal pillar's stress state after roof-cutting can be predicted, providing more accurate load boundary conditions for concrete wall reinforcement. When the key basic roof strata are cut off, the overlying strata lose their continuous bearing capacity, and the stress concentration factor... Significantly reduced, according to the formula for calculating the boundary of the plastic zone. It can be seen that, with As the value decreases, the range of the plastic zone x0 shrinks accordingly. Simultaneously, disrupting the continuity of the rock strata reduces the transmission of lateral stress and improves the stress environment of adjacent roadways. This fundamentally solves the problem of deformation and instability of the surrounding rock in adjacent roadways on the goaf side of a section coal pillar under deep-well mining conditions.

[0035] Furthermore, in S4, the fractured zone is used as the influence range for calculating the surrounding rock pressure. The static load q exerted by the surrounding rock pressure on the concrete wall is calculated, and the static load q is corrected using a dynamic load factor to obtain the support load Q borne by the concrete wall. The correction of the static load using a dynamic load factor to obtain the support load Q includes: calculating the static load q borne by the concrete wall based on the fractured zone range and geological parameters. Where: q represents the concrete wall support load; x represents the width of the concrete wall; The distance from the outer edge of the concrete wall support to the roof cantilever is shown; γ represents the unit weight of the separated rock block in the roof; h represents the mining height; θ represents the shear angle. H represents the coal seam dip angle; H represents the roof collapse height; the dynamic load coefficient η is set to determine the dynamic pressure exerted on the immediate roof by the rotation of the basic roof stratum; 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] Among them, the surrounding rock pressure refers to the pressure load exerted on the concrete wall by the unstable surrounding rock within the fractured zone, including the self-weight pressure of the fractured coal and the additional pressure transmitted to the concrete wall by the overlying strata through the fractured zone. This pressure is the main basis for the load-bearing design of the concrete wall, reflecting the load requirements of the artificial support structure on the surrounding rock that has lost its self-bearing capacity.

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

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

[0039] Roof separation rock blocks refer to rock mass segments that separate from the intact roof due to the fracturing and loss of support of the underlying coal seam. These rock blocks have lost their original load-bearing foundation, and their weight needs to be borne by artificial support structures such as concrete walls, making them a major component of the surrounding rock pressure.

[0040] The roof collapse height H refers to the total height of the collapsed rock strata from the coal seam roof upwards within the goaf, measured in meters (m). This height reflects the degree of collapse of the rock strata above the goaf, influencing the stress transmission path and load magnitude from the overlying rock strata to the concrete wall, and is an important parameter for calculating surrounding rock pressure.

[0041] The rotation of the basic roof strata refers to the fracturing of the basic roof strata after mining at the working face, and the rotational movement of the strata towards the goaf with the fracturing point as the fulcrum. This rotation exerts a squeezing and impacting effect on the underlying immediate roof strata, and is the main mechanism for generating dynamic load effects.

[0042] The dynamic load factor η refers to the correction factor 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 strata. It is a dimensionless parameter with a value range of 1-3.

[0043] Compared to existing technologies, the advantages of this application are:

[0044] (1) By establishing the vertical stress distribution at each point in the plastic zone during the coal seam loading process, and combining 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 fracture zone, crack zone, and original structure zone of the coal pillar are scientifically divided. Combined with the roof cutting and roadway protection technology in step S3, the basic roof strata at the boundary between the fracture zone and the caving zone are cut by hydraulic fracturing, thereby destroying the continuity and integrity of the overlying strata and reducing the lateral constraint and pressure transmission of the overlying strata on the coal pillar. When the overlying strata lose their 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, with the stress concentration factor The reduction in the plastic zone range The width of the coal pillar is reduced accordingly, thus effectively controlling the damage width.

[0045] (2) The optimal target stratum for roof cutting and borehole depth are determined through rock structure analysis. The load requirements for concrete wall support are accurately calculated by combining the rock block separation method and dynamic load coefficient correction. This organically combines roof cutting and roadway protection technology with concrete wall reinforcement technology. Compared with the independent application of the two technologies in the prior art, this application can reduce the damage width of the coal pillar 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 combination of roof cutting and concrete wall reinforcement forms a mechanical coupling mechanism, blocking stress transmission paths from two dimensions. First, roof cutting disrupts the continuity of the target stratum, causing the horizontal stress that was originally transmitted along the stratum to concentrate and release at the roof cutting location, thus blocking the direct transmission path of far-field stress to adjacent roadways. Second, the installation of the concrete wall changes the near-field stress distribution pattern. Based on the wall thickness and layout 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 greater than that of broken coal, stress is preferentially transmitted through the concrete wall, reducing the stress component transmitted to adjacent roadways through the coal pillar. The synergistic effect of the two technologies effectively disperses and absorbs lateral mining stress during transmission. Attached Figure Description

[0047] This application will be further described by way of exemplary embodiments, which will be described in detail with reference to the accompanying drawings. These embodiments are not limiting; in these embodiments, the same reference numerals denote the same structures, wherein:

[0048] Figure 1 This is an exemplary flowchart illustrating a method for reinforcing the concrete wall on the goaf side of a mining section according to some embodiments of this application;

[0049] Figure 2 This is a schematic diagram of the roof rock strata structure according to some embodiments of this application;

[0050] Figure 3 This is a schematic diagram of a coal seam structure along the goaf, based on some embodiments of this application;

[0051] Figure 4 This is a schematic diagram of the top plate structure and vertical stress distribution according to some embodiments of this application;

[0052] Figure 5 This is a process flow diagram of constructing a concrete wall using the spraying process shown in some embodiments of this application. Detailed Implementation

[0053] The methods and systems provided in the embodiments of this application will now be described in detail with reference to the accompanying drawings.

[0054] The minefield covers an area of ​​10.8186 km², with a production scale of 1.2 million tons per year. Mining depths range from 900 meters to 300 meters above sea level. The mine employs a central parallel ventilation system. The mine is divided into two levels: Level 1 mines the No. 3 coal seam, and Level 2 mines the No. 15 coal seam. Currently, the mine is mining the No. 3 coal seam at Level 1, with an elevation of +663m. It is divided into three panels, namely, Panel 1, Panel 2, and Panel 3, from north to south within the minefield. The first mining area is the central part of Panel 1.

[0055] According to the geological description of the first panel area and the mining status data provided by the mine, the mine is currently carrying out mining activities in the first panel area (east). The first panel area (east) is mined on two wings. The west wing is equipped with working faces 3101, 3103, 3105, 3107, and 3109, and the east wing is equipped with working faces 3102, 3104, 3106, 3108, and 3110.

[0056] The coal seam in the No. 1 panel of Xiaoxi Coal Mine is buried at a depth of approximately 600-700 meters, a typical deep-well mining operation. It employs a one-time full-height mining (high-height mining), single-wing sequential mining, and double-roadway excavation, with a coal pillar width of 25 meters. The mining of this face has a strong impact on adjacent roadways; for example, during the mining of the 3108 longwall face, the deformation of the gas control roadway in the east wing was relatively severe.

[0057] The immediate roof of the tunnel is mostly fine-grained sandstone and sandy mudstone, with some areas consisting of silty mudstone, mudstone, or silty mudstone. The rock mass has low strength and is easily weathered and broken. The tunnel has a large cross-section, making support difficult.

[0058] According to the "Test Report on Physical and Mechanical Parameters of Roof Rock Strata in No. 3 Coal Seam of Yangcheng Yangtai Group Xiaoxi Coal Industry Co., Ltd.", the uniaxial compressive strength of the coal is 10.4-14.5 MPa. The coal body has low strength. After the working face is mined, the coal pillars in the roadway will undergo plastic deformation due to the support pressure.

[0059] The complex structure of the block is complex and has poor integrity, resulting in weakened self-bearing capacity. The surrounding rock of the roadway exhibits loose and fractured characteristics, causing the anchor bolts (cables) to lose their stable support foundation and fail to play an active support role. Under the combined action of rock mass swelling pressure and deformation pressure, they flow and are squeezed into the roadway, forming net-like and falling debris phenomena. As the degree and depth of surrounding rock fracture continue to increase, the anchor bolts (cables) gradually lose their support capacity, and the deformation and instability of the roadway intensifies.

[0060] Given the above reasons, although various comprehensive measures were taken to manage the gas in the east wing of the No. 1 panel gas control roadway, significant deformation still occurred (large roof and floor displacement, sidewall convergence, and support failure, etc.) despite increased costs. Traditional coal pillar support and roadway reinforcement technologies have limited effectiveness. Therefore, under the original comprehensive management approach, research is proposed on the mechanical coupling mechanism and synergistic bearing technology of reinforced concrete walls on the goaf side of the section coal pillar to prevent severe deformation of adjacent roadways caused by subsequent working face mining disturbances. Figure 1 As shown, the geological parameters of the coal pillar in the collection section are obtained. The coal pillar in the collection section is the coal body between the goaf formed by the working face and the area to be mined. Based on the geological parameters of the coal pillar, the boundary position of the plastic zone is calculated. According to the location of the plastic zone boundary The coal pillar section is divided into a fractured zone, a ruptured zone, and a zone with original structure. Based on the geological parameters of the coal pillar, the basic roof strata are determined, and the bearing capacity of each basic roof strata is calculated. The basic roof stratum with the lowest bearing capacity is selected as the target stratum for roof cutting, and the hydraulic fracturing borehole depth required to penetrate the target stratum is determined. The fractured zone is used as the influence range for calculating the surrounding rock pressure. The static load q exerted by the surrounding rock pressure on the concrete wall is calculated, and the static load q is corrected using a dynamic load factor to obtain the support load Q. Based on the support load Q, the concrete strength grade of the concrete wall, the thickness of the concrete wall in the roadway cross-section perpendicular to the roadway axis, and the horizontal distance between the top of the concrete wall and the edge of the goaf side roof are determined.

[0061] Specifically, during coal seam mining, the roof strata are disturbed, and due to changes in the supporting structure between the strata, they undergo different stress field evolutions and rebalancing processes. The roof strata at different locations within the disturbed area will experience varying degrees of damage, exhibiting different structural characteristics, such as... Figure 2 As shown, it forms three horizontal zones (A-coal seam support zone, B-separation fracture zone, C-fractured collapse zone) and three vertical zones (Ⅰ-caving zone, Ⅱ-fracture zone, Ⅲ-bending subsidence zone).

[0062] After undergoing the evolution of support pressure, coal seams along the goaf edge often exhibit a fractured-fractured-original structural characteristic from the edge to the interior (e.g. Figure 3(As shown). The edge fractures and internal fractures are caused by the strength failure during the redistribution of support pressure formed during the mining process. The size of the fracture zone and fracture zone is affected by factors such as coal properties, rock strata structure, mining thickness, goaf size, and mining depth.

[0063] For longwall mining faces, yield line analysis is used to deduce that the roof yield failure line will form an "O-X" fracture in the early stages of mining. During subsequent mining, internal and external stress fields will form in front of and laterally of the working face. The lateral roof structure and vertical stress distribution of the working face are shown in the following figures. Figure 4 As shown.

[0064] During the loading process, coal seams develop plastic zones, and the internal stress field is usually distributed within these zones. Considering coal seam compression, the expression for the vertical stress at each point in the plastic zone is:

[0065] (1);

[0066] In the formula: This refers to the vertical stress at each point in the plastic zone; denoted as uniaxial compressive strength of the coal seam; C is the cohesion between the coal seam and the roof and floor; f is the coal body firmness coefficient; h is the mining thickness of the coal seam. θ is the coal seam compression angle; x is the distance into the coal face. The distance into the coal face where the maximum supporting pressure is measured.

[0067] Before the roof of the goaf fractures, the support pressure at the junction of the elastic-plastic zones reaches its maximum value, i.e., when... hour,

[0068] (2);

[0069] in: The maximum stress concentration factor of the coal seam can be used as the criterion. Substituting this condition into equation (1) will yield the range of the plastic zone. The expression is:

[0070] (3);

[0071] in: The coefficient relating the maximum bearing pressure to the coal seam strength condition is expressed as follows:

[0072] (4);

[0073] It can be seen that the burial depth H and the maximum stress concentration factor The larger the value, the larger the range of the plastic zone. The smaller, The larger.

[0074] Based on the above analysis, under certain mining depth and given coal seam conditions, the range of the plastic zone is directly proportional to the coal seam mining thickness h. The specific boundary of the internal stress field is determined by the strength of the rock beam and the clamping force required for fracture. The larger the mass and strength of the rock beam, 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-plastic zone of the coal pillar. At this stage, the roadway will be affected by the lateral stress, resulting in large deformation and roadway cross-section shrinkage.

[0075] Therefore, in order to alleviate the deformation of the goaf roadway, research should be carried out on the technology of roof cutting and roadway protection, as well as the mechanical coupling and synergistic bearing technology of the reinforced concrete wall on the goaf side of the coal pillar in the mining roadway. The aim 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 adjacent roadways, and improve the stress environment of adjacent roadways.

[0076] Hydraulic fracturing roof cutting and tunnel protection calculations, required roof cutting height at the working face:

[0077] Based on the distribution characteristics of the No. 3 coal seam in Xiaoxi Coal Mine:

[0078] ;

[0079] The basic top rock layer entering the fracture zone can be calculated using the following formula:

[0080] , The thickness of the i-th basic rock layer from bottom to top; M represents the thickness of the i-th basic top layer from bottom to top; M- represents the coal seam mining height. The rock fragmentation coefficient of the basic top and its additional rock layers; ρ is the rock fragmentation coefficient of the immediate roof and its adjacent strata; h is the thickness of the immediate roof;

[0081] Substituting the relevant data from the working face into the above formula, we can deduce that the third basic top rock beam, i.e., the rock layer above the 16.84m sandy mudstone, is a fracture zone rock layer, and the rock layer below it is a caving zone rock layer. Therefore, the vertical depth of the fracturing borehole should reach above the third basic top rock beam. To ensure the fracturing effect, the fracturing location should be as deep as possible, penetrating half the thickness of the fourth layer of fine-grained sandstone.

[0082] In summary, the required cut-off height along the air gap is 40m.

[0083] Based on the geological and mining conditions of the Xiaoxi Coal Mine working face, the "separated rock block method" was chosen to calculate the surrounding rock pressure on the concrete wall. This method is a theoretical approach to calculating support loads based on the surrounding rock instability mechanism, used to determine the magnitude of the surrounding rock pressure that the concrete wall needs to bear. The core principle of this method is to treat the rock mass separated from the intact surrounding rock within a certain influence range above the concrete wall due to the fracturing of the underlying coal seam as an independent load unit for analysis and calculation. Specifically, the separated rock block method assumes that when the coal pillar fracture zone loses its bearing capacity, the roof strata above it will separate and fracture under the action of gravity and overlying pressure, forming relatively independent rock block structures. The weight of these separated rock blocks and the pressure from the overlying strata transmitted through them constitute the main static load borne by the concrete wall. This method calculates the surrounding rock pressure acting on the concrete wall by establishing a geometric model of the separated rock blocks and combining parameters such as the block's unit weight, thickness, and distribution range, using the principle of mechanical equilibrium.

[0084] 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 mine pressure of the concrete wall is as follows:

[0085] ;

[0086] In the formula: —Concrete wall support load;

[0087] —The width of the concrete wall is taken as 3m;

[0088] —The overhang distance on the outer side of the concrete wall support is 1.5m;

[0089] —The unit weight of the separated rock block in the top plate is taken as 27kN / m3;

[0090] —Mining height, 6.2m;

[0091] —The shear angle is selected as 26° based on experience;

[0092] —The coal seam dip angle is 2°;

[0093] —The collapse height of the roof slab is 31m, with the rock strata fragmentation coefficient taken as 1.2.

[0094] Calculations show that when the support width is 3m, the pressure borne by the roadside support is:

[0095] q = 6.5 MPa;

[0096] The load on the concrete wall originates from the weight of the separated rock blocks from the immediate roof and the dynamic pressure exerted on the immediate roof by the rotation of the main roof. The dynamic pressure exerted on the immediate roof by the rotation of the main roof has been difficult to determine; the traditional approach is to measure the so-called dynamic load coefficient η using actual measurements. Based on fully mechanized mining experience, η is generally 1–2. When designing the concrete wall support, we take 2. Thus, the maximum load on the concrete wall is:

[0097] 2 =2×6,5=13MPa;

[0098] In summary, the concrete strength grade C30 used for the goaf side concrete wall support of the Xiaoxi Coal Mine section is sufficient to meet the strength requirements.

[0099] Pile-spraying technology is an innovative method for constructing concrete walls. Sand, cement, and JCT-1 type sprayable concrete additive are thoroughly mixed in a specific ratio in a humid (or dry) environment and then placed into a spraying device. Compressed air is used to propel the mixture through pipes to the nozzles, where it combines with water and is sprayed at high speed onto the construction surface to form a coating or wall. The specific process is as follows: Figure 5 As shown. The spraying process for constructing concrete walls or sprayed layers has the following characteristics: The construction process only requires temporary baffles; a single spray application can achieve a relatively thick layer; the wall surface forms quickly, eliminating the need for support frames and subsequent maintenance; the sprayed surface has strong adhesion, effectively suppressing dust and sealing moisture, with a rebound rate of less than 5%; initial hardening is rapid, maximum hardness does not decrease, and the final strength reaches C30 or higher; it has good grafting performance, with the sprayed layer and the rock layer sharing the weight; the wall has high density, high strength, micro-expansion, a dense texture, and is airtight; the construction machinery used is small in size, quick and convenient to operate, reducing costs and labor requirements, and the construction speed is relatively fast.

[0100] The use of shotcrete technology to construct retaining walls is not only fast and simple to implement, but also ensures the strength of the concrete, offering significant cost advantages and fully complying with safety standards for underground coal mining operations. The construction speed of shotcrete walls primarily depends 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. It plays a crucial role in increasing wall construction speed, reducing labor intensity, and shortening construction time, and is essential for promoting the safe and efficient operation of coal mines.

[0101] The foregoing illustrative description of the present application and its embodiments is not restrictive and can be implemented in other specific forms without departing from the spirit or essential characteristics of the present application. The accompanying drawings are only one embodiment of the present application, and the actual structure is not limited thereto. Therefore, if those skilled in the art are inspired by this description and design similar structures and embodiments without departing from the spirit of the present application, such designs should fall within the scope of protection of this application. Furthermore, the word "comprising" does not exclude other elements or steps, and the word "a" preceding an element does not exclude the inclusion of "a plurality" of that element. Terms such as "first," "second," etc., are used to indicate names and do not indicate any specific order.

Claims

1. A method for reinforcing the concrete wall on the goaf side of a mining section, characterized in that, include: Geological parameters of coal pillars in the collection section are collected. The coal pillars are the coal bodies between the goaf formed by the working face and the area to be mined. Calculate the location of the plastic zone boundary based on the geological parameters of the coal pillar. According to the location of the plastic zone boundary The coal pillar section is divided into a fractured zone, a cracked zone, and a zone with original structure. Based on the geological parameters of the coal pillar, the basic roof strata are determined, the bearing capacity of each basic roof strata is calculated, the basic roof strata with the smallest bearing capacity are taken as the target stratum for cutting the roof, and the depth of the hydraulic fracturing borehole required to penetrate the target stratum is determined. The fractured zone is taken as the influence range for the calculation of surrounding rock pressure. 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 factor, and the support load Q is obtained. Based on the support load Q, determine the concrete strength grade of the concrete wall, the thickness of the concrete wall in the roadway cross-section perpendicular to the roadway axis, and the horizontal distance between the top of the concrete wall and the edge of the goaf side roof.

2. The method for reinforcing the concrete wall on the goaf side of the mining section according to claim 1, characterized in that: Based on the location of the plastic zone boundary The coal pillar section is divided into a fractured zone, a fractured zone, and a zone with original structure, including: Starting from the edge of the goaf, the coal pillar is divided into three zones along the horizontal direction of the coal pillar: The fractured zone, which is a region from the edge of the goaf to the boundary of the fractured zone, is a region where the coal body is completely broken and has lost its bearing capacity. From the boundary between the fractured zone and the plastic zone The fracture zone within which the coal body undergoes plastic deformation but still retains some load-bearing capacity; From the boundary of the plastic zone The original structural zone within the coal pillar, which maintains the original structure and complete load-bearing capacity of the coal body.

3. The method for reinforcing the concrete wall on the goaf side of the mining section according to claim 1, characterized in that: Determining the required hydraulic fracturing cut-out depth to penetrate the target stratum includes: Based on the thickness of each rock stratum, the rock strata structure is identified in ascending order from the coal seam, with the first layer above the coal seam having a thickness less than a threshold. And the uniaxial compressive strength is below the threshold. The rock strata are taken as the immediate top, and the thickness above the immediate top is greater than the threshold. And the uniaxial compressive strength is greater than the threshold. The rock strata are sequentially designated as the i-th basic top rock stratum; Based on the identified basic top rock layers and their coefficients of fragmentation Calculate the critical thickness of the i-th basic top rock layer entering the fracture zone from bottom to top. ; Based on the calculated critical thickness of each basic top rock layer By comparing the actual rock strata thickness with the actual samples, the highest basic top rock stratum entering the fracture zone is determined. The part above the highest basic top rock stratum is taken as the fracture zone rock stratum, and the part below it is taken as the caving zone rock stratum. 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 depth of the basic borehole is determined according to the vertical distance from the coal seam roof to the target layer for cutting the top. The depth of the hydraulic fracturing top-cutting borehole is calculated based on the basic borehole depth.

4. The method for reinforcing the concrete wall on the goaf side of the mining section according to claim 3, characterized in that: Calculate the critical thickness of the i-th basic top rock layer entering the fracture zone from bottom to top. The following formula is used: ;in, The critical thickness for the i-th basic top rock layer to enter the fracture zone from bottom to top; M represents the thickness of the i-th basic top rock layer from bottom to top; M represents the coal seam mining height. The coefficient of rock fragmentation of the basic top stratum; The coefficient of rock fragmentation of the immediate overlying stratum; The thickness is the direct top thickness.

5. The method for reinforcing the concrete wall on the goaf side of the mining section according to claim 3, characterized in that: The hydraulic fracturing top-cutting borehole depth is calculated based on the basic borehole depth, including: A penetration distance is set above the basic drilling depth, and the penetration distance is 1 / N of the thickness of the adjacent rock layer above the target layer. The sum of the basic borehole depth and the penetration distance is taken as the hydraulic fracturing top-cutting borehole depth, so that hydraulic fracturing can completely penetrate the top-cutting target layer and form a fracture network.

6. The method for reinforcing the concrete wall on the goaf side of the mining section according to claim 5, characterized in that: The value of N is either 2 or 3.

7. The method for reinforcing the concrete wall on the goaf side of the mining section according to claim 5, characterized in that: Obtain the support load Q borne by the concrete wall; The static load is corrected using a dynamic load factor to obtain the support load Q, which includes: Based on the extent of the fractured zone and geological parameters, calculate the static load q borne by the concrete wall: ； Where: q represents the concrete wall support load; x represents the width of the concrete wall; Indicates the overhang distance on the outer side of the concrete wall support; The unit weight of the separated rock block in the roof is represented by h; the mining height is represented by θ; and the shear angle is represented by θ. H represents the dip angle of the coal seam; H represents the roof collapse height. Set the dynamic load coefficient η for the dynamic pressure exerted on the immediate top layer by the rotation of the basic top layer; The support load Q borne by the concrete wall is obtained based on the dynamic load factor η and the static load q.

8. The method for reinforcing the concrete wall on the goaf side of the mining section according to claim 7, characterized in that: The dynamic load factor η ranges from 1 to 3.

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