Strong mining roadway thick-layer transboundary supporting method

Abstract

The invention relates to a strong mining roadway thick-layer transboundary supporting method, and belongs to the technical field of roadway supporting, and the method comprises the steps that geometric parameters, rock mass mechanical parameters and stress field parameters of a strong mining roadway are obtained; based on the geometric parameters, the rock mass mechanical parameters and the stress field parameters, the deformation process and the damage process of the strong mining roadway surrounding rock are simulated, and the incidence relation between each deformation stage of the strong mining roadway surrounding rock and the roof partition damage characteristic is obtained; determining a key stage of irreversible deformation of the strong mining roadway surrounding rock based on the association relationship; determining the critical time of transboundary anchoring of the thick layer based on the key stage; determining the support parameters of the anchor rod and the critical anchoring length of the thick layer transboundary; and supporting the top plate of the strong mining roadway based on the critical time, the supporting parameters and the thick-layer transboundary anchoring length. The method has the effect of improving the stability of the strong mining roadway surrounding rock.

Description

Technical Field

[0001] The present application relates to the technical field of tunnel support, and in particular to a method for cross-boundary support of thick layers in a strongly mining tunnel. Background Art

[0002] During high-intensity coal mining, especially in complex conditions such as extremely thick coal seams, gob-side entry retention, gob-side entry driving, and isolated island working faces, roadways are often subject to intense mining activity. This can lead to significant deformation of the surrounding rock, severe roof subsidence, and failure of support structures. These roadways are known as "high-intensity mining roadways." These surrounding rock faces are subject to high stress and strong disturbance, resulting in poor stability and prone to local or overall instability, seriously impacting mine safety and production.

[0003] Existing data indicate that surrounding rock failure in heavily mined roadways exhibits zonal characteristics, including fractured, plastic, and elastic zones. The distribution of these zones is influenced by in-situ rock stress, mining stress, and abutment pressure. Field tests and numerical simulations demonstrate that the roadway roof failure zone and tunneling-influenced zone are extensive and complex. In particular, when key lateral roof blocks in the goaf fracture and rotate and sink, the roadway is subjected to intense dynamic loads, further exacerbating surrounding rock failure.

[0004] Currently, surrounding rock control for highly dynamic roadways is primarily focused on three aspects: first, optimizing mining layout and coal pillar size to improve the surrounding rock stress environment; second, employing active pressure-relieving technologies such as hydraulic fracturing and deep-hole blasting to reduce stress concentration; and third, strengthening support structures, particularly bolt-and-cable combined support systems, to increase the surrounding rock's self-bearing capacity. Bolt-and-cable support is widely used in engineering projects because it can provide timely support resistance, inhibit delamination, and control deformation.

[0005] In recent years, researchers have developed new support methods, such as constant-resistance, high-deformation anchor cables and deep, stable rock anchoring, to accommodate the large deformations and high stresses experienced under intense mining conditions. However, current rock control methods using combined anchor and cable support do not fully meet the requirements for maintaining and controlling tunnels subjected to intense mining conditions, and large surrounding rock deformation and severe dynamic rock pressure still occur in these tunnels. Summary of the Invention

[0006] In order to improve the stability of the surrounding rock of a roadway subjected to strong mining, this application provides a thick layer cross-boundary support method for a roadway subjected to strong mining, which adopts the following technical solutions:

[0007] A method for cross-boundary support of thick layers in a strong mining tunnel, comprising:

[0008] Obtain geometric parameters, rock mass mechanical parameters and stress field parameters of strong mining tunnels;

[0009] The deformation and failure processes of the surrounding rock of the strong mining tunnel are simulated based on the geometric parameters, rock mass mechanics parameters and stress field parameters, and the correlation between each deformation stage of the surrounding rock of the strong mining tunnel and the failure characteristics of the roof partition is obtained;

[0010] Determine a critical stage in which irreversible deformation of the surrounding rock of the strongly mining roadway occurs based on the correlation relationship, wherein the critical stage is a stage in the deformation process of the surrounding rock of the strongly mining roadway;

[0011] Determining the critical time for thick layer cross-boundary anchoring based on the key stages;

[0012] Determining anchor support parameters and a critical anchoring length for thick-layer crossings, wherein the critical anchoring length is the sum of the damaged layer of the roof rock mass caused by the mining and excavation process of the surrounding rock working face of the highly active mining roadway and the anchoring boundary;

[0013] The roof of the heavily mined tunnel is supported based on the critical time, support parameters and thick layer cross-boundary anchoring length.

[0014] By adopting the above technical scheme, the deformation process and destruction process of the surrounding rock of the strong mining tunnel are simulated to determine the correlation between the deformation stage and the roof partition destruction characteristics, thereby determining the key stage of irreversible deformation through the roof partition destruction characteristics, and then determining the critical time of thick layer cross-border anchoring, ensuring that the support measures are implemented at the optimal time node, and effectively avoiding the safety hazards caused by premature failure of the surrounding rock or delayed support. The critical anchoring length is determined by the damage depth of the roof rock mass and the stable anchoring edge, so that the anchor end of the anchor cable penetrates into the stable rock layer that is not affected by mining, forming a high-strength bearing structure spanning the broken zone and the plastic zone, thereby improving the adaptability to large deformation of the surrounding rock and severe mine pressure, and to a certain extent improving the stability of the surrounding rock of the strong mining tunnel.

[0015] Optionally, the deformation stages of the surrounding rock of the intense mining tunnel include an instantaneous stage, a sudden change stage, a slow increase stage and a destruction stage.

[0016] Optionally, the roof zoning destruction characteristics of the surrounding rock of the strong mining tunnel are as follows: the roof rock layers show coordinated deformation characteristics, uncoordinated deformation characteristics, delamination characteristics and three-hinge arch structure characteristics from shallow to deep.

[0017] Optionally, the relationship between each stage and the roof partition failure characteristics is:

[0018] The instantaneous stage corresponds to the coordinated deformation characteristics; the drastic change stage corresponds to the uncoordinated deformation characteristics; the slow increase stage corresponds to the delamination characteristics; and the destruction stage corresponds to the three-hinged arch structure characteristics.

[0019] Optionally, determining the critical time for thick layer cross-boundary anchoring based on the key stage includes:

[0020] Determining the stage where the surrounding rock of the intense mining tunnel produces irreversible deformation based on the roof zoning damage characteristics;

[0021] The starting time of the irreversible deformation stage is taken as the critical time;

[0022] The irreversible deformation stage is one of the deformation stages of the surrounding rock of the intense mining tunnel.

[0023] By adopting the above technical solution, by identifying the starting moment of irreversible deformation of the surrounding rock as the critical anchoring time, it is ensured that support measures are intervened in time before the surrounding rock structure is completely destroyed, effectively preventing further expansion of deformation.

[0024] Optionally, determine the anchor support parameters, including:

[0025] Calculate the shear stress on each longitudinal section parallel to the neutral layer when the roof rock beam undergoes pure bending deformation;

[0026] determining a shear force applied to each anchor rod based on the shear stress;

[0027] Obtain the stiffness of the anchor rod and the increment of the anchor rod length;

[0028] Calculating the tension borne by each anchor rod based on the stiffness and the increment of the anchor rod length;

[0029] The support parameters of the anchor rod are determined based on the shear force and the tensile force.

[0030] Optionally, determine the critical anchorage length of the anchor bolt across the thick layer, including:

[0031] The mining and excavation process of the surrounding rock working face of the strong mining tunnel was simulated to obtain the damaged layer and anchoring boundary of the roof rock mass;

[0032] Calculating the sum of the damaged layer and the anchoring boundary;

[0033] The sum is taken as the critical anchoring length.

[0034] By adopting the above technical solution, by extending the anchoring length to beyond the stable anchoring boundary below the damaged layer, the anchoring end of the anchor cable can be effectively avoided from being in the crushing or plastic deformation zone, thereby improving the overall bearing capacity of the anchoring system and enhancing the resistance to large deformation of the surrounding rock and severe mine pressure.

[0035] Optionally, the anchor rod is a flexible anchor rod.

[0036] Optionally, also include:

[0037] Obtain the formula for the maximum tensile stress when the roof rock beam undergoes pure bending deformation;

[0038] The relationship between the maximum compressive stress and thickness is obtained based on the moment of inertia of the cross section about the neutral axis and the maximum tensile stress formula;

[0039] The relationship is that the maximum tensile stress is inversely proportional to the cube of the thickness, that is, the thicker the anchoring layer, the stronger the bearing capacity of the roof rock beam.

[0040] Optionally, based on the entire process of deformation and destruction of the strong mining tunnel, it is determined that the formation and upward extension of longitudinal cracks play a dominant role in the formation of the roof partition failure structure, wherein the formation and upward extension of longitudinal cracks occur after the instantaneous stage.

[0041] In summary, this application has the following beneficial technical effects:

[0042] By identifying the starting moment of the surrounding rock entering irreversible deformation as the critical anchoring time, it is ensured that the support measures are timely intervened before the surrounding rock structure is completely destroyed, effectively preventing the deformation from further expanding. When the anchor end is arranged at a depth greater than the critical anchoring length, the tensile stress in the shallow rock mass will be transferred to the deep overburden. The shallow rock mass at the critical anchoring length is sensitive to tensile stress, while the surrounding rock at the deep end is stable and has strong bearing capacity, and is not very sensitive to tensile stress. Therefore, the application of cross-border support can play a dual role of stress equalization and stress dilution, avoiding the shallow damaged rock mass from being directly subjected to tensile stress, and effectively controlling the deformation of the shallow surrounding rock; and the thick cross-layer The boundary support provides two-way constraint to the rock mass between the anchor end and the tail of the anchor rod, so that it is in a three-way compressive state, strengthening the bearing capacity of the surrounding rock and preventing further expansion of rock damage; on the other hand, it is conducive to the two-way continuous transmission of stress in the anchor body, so that it has a direct connection with the deep overburden, and mobilizes a larger range of rock mass to participate in the bearing, realizes the homogenization of stress and the deep and shallow linkage of displacement, and further limits the displacement of shallow rock mass; at the same time, under the action of thick cross-boundary anchoring, the tensile stress is transferred to the deep surrounding rock, which leads to the critical anchoring length in the shallow range, the strain zero point gradually moves downward, the surrounding rock deformation range is reduced, and the deformation of the surrounding rock is effectively controlled. BRIEF DESCRIPTION OF THE DRAWINGS

[0043] Figure 1 It is a flow chart of a method for cross-boundary support of thick layers in a strongly mining tunnel in an embodiment of the present application.

[0044] Figure 2 It is a schematic diagram showing the plate partition destruction characteristics in the embodiment of the present application.

[0045] Figure 3 It is a time displacement diagram reflecting the entire process of deformation and destruction of surrounding rock in a strong mining tunnel in the embodiment of the present application.

[0046] Figure 4It is a schematic diagram of the shear stress at any point k on the cross section at a distance y from the neutral axis when the roof rock beam is bent in the embodiment of the present application.

[0047] Figure 5 This is a schematic diagram showing the top rock beam bearing shear force in an embodiment of the present application.

[0048] Figure 6 This is a schematic diagram showing the shear force borne by the anchor rod when there is an anchor rod in the embodiment of the present application.

[0049] Figure 7 This is a schematic diagram showing the tension borne by the anchor rod when the rock layer is delaminated in the embodiment of the present application.

[0050] Figure 8 Schematic diagram showing the positions of the EIZ, EDZ and plastic boundary in the stress-strain curve in the embodiment of the present application.

[0051] Figure 9 It is a schematic diagram showing the distribution of EDZ, EIZ, Intact Zone and plastic zone of surrounding rock in the embodiment of the present application.

[0052] Figure 10 It is a schematic diagram of the coal mine mining layout in the embodiment of the present application.

[0053] FIG11 is a schematic diagram showing the distribution of anchor rods of the original support in an embodiment of the present application, (a) is a side view, and (b) is a top view.

[0054] FIG12 is a schematic diagram of the distribution of anchor rods for thick-layer cross-border anchoring support in an embodiment of the present application, (a) is a side view, and (b) is a top view.

[0055] Figure 13 This is a stress field diagram reflecting the support in the embodiment of the present application, (a) is no support, (b) is the original support, and (c) is the thick layer cross-border anchor support.

[0056] Figure 14 It is a curve diagram reflecting the simulation results of the plate partition failure characteristics in the embodiment of the present application, (a) is no support, (b) is the original support, and (c) is the thick layer cross-border anchor support.

[0057] FIG15 is a graph showing the roof subsidence of five measuring stations in the embodiment of the present application, (a) and (b) are 14 -1 (c) is a bar graph showing the roof deformation measurement of the air inlet chute at the 105 working face, and (c) is a bar graph showing the three stages of roof subsidence.

[0058] Figure 16 This is a diagram showing the internal rock structure of the top plate of borehole 1# in the embodiment of the present application.

[0059] FIG17 is a mechanical principle diagram showing the thick layer cross-boundary anchoring effect in an embodiment of the present application, including (a) tensile stress driving effect and (b) compressive stress continuous transmission effect. DETAILED DESCRIPTION

[0060] The present application is further described in detail below with reference to the accompanying drawings.

[0061] This specific embodiment is merely an explanation of the present application and is not a limitation of the present application. After reading this specification, those skilled in the art may make non-creative modifications to the present embodiment as needed, but as long as they are within the scope of the claims of the present application, they are protected by the patent law.

[0062] To make the purpose, technical solutions, and advantages of the embodiments of this application more clear, the technical solutions in the embodiments of this application will be clearly and completely described below in conjunction with the drawings in the embodiments of this application. Obviously, the described embodiments are part of the embodiments of this application, not all of the embodiments. Based on the embodiments in this application, all other embodiments obtained by ordinary technicians in this field without making creative efforts are within the scope of protection of this application.

[0063] In this document, the term "and / or" simply describes a relationship between related objects, indicating that three possible relationships exist. For example, "A and / or B" can represent: A exists alone, A and B exist simultaneously, or B exists alone. Furthermore, the character " / " in this document, unless otherwise specified, generally indicates an "or" relationship between the related objects.

[0064] The embodiments of the present application are described in further detail below with reference to the accompanying drawings.

[0065] like Figure 1 As shown, a method for cross-boundary support of thick layers in a roadway with strong mining is described as follows (steps S101 to S106):

[0066] Step S101, obtaining geometric parameters, rock mass mechanical parameters and stress field parameters of a strong mining tunnel;

[0067] In this embodiment, the geometric parameters of the intensive mining tunnel are obtained by methods including but not limited to on-site measurements, construction drawings and design data, and geological modeling software.

[0068] Among them, on-site measurement: tunnels are measured using total stations, laser scanners, 3D geological radars and other equipment; construction drawings and design data: mine design drawings contain the geometric parameters of the tunnels, including tunnel boring machine construction logs and CAD drawings; modeling software: such as Surpac, MineSight, Micromine, etc., can be used to construct 3D models of tunnels and extract geometric parameters.

[0069] Among them, geometric parameters include but are not limited to the length, width and height of the strong mining tunnel mining, rock mechanical parameters include but are not limited to the physical and mechanical properties of the rock mass of the roof, stress field parameters are the core indicators of the internal force distribution state in the rock mass or structure, and stress field parameters include but are not limited to original rock stress and mining stress.

[0070] Step S102: simulating the deformation and failure processes of the surrounding rock of the heavily mining roadway based on geometric parameters, rock mass mechanics parameters, and stress field parameters to obtain the correlation between each deformation stage of the surrounding rock of the heavily mining roadway and the failure characteristics of the roof partition;

[0071] An on-site investigation of the mining process of the strong mining tunnel found that the deformation process of the surrounding rock of the strong mining tunnel has obvious stages, among which the deformation stages include instantaneous stage, drastic change stage, slow increase stage and destruction stage, and the roof in each deformation stage also shows different structural characteristics.

[0072] Instantaneous stage: In the initial stage of surrounding rock deformation, the rock mass that plays a dominant role in the development of surrounding rock deformation usually has elastic deformation characteristics. In this stage, the surface of the tunnel roof is flat, the internal structure of the rock layer is not damaged, the bearing state is good, and the rock mass is still in an elastic state; Sudden change stage and slow increase stage: In the middle stage of surrounding rock deformation, the rock mass that plays a dominant role in the development of surrounding rock deformation usually has shear fragmentation characteristics. In this stage, the surface of the roof begins to break, the rock blocks fall in the shape of a "net bag", and the shallow surrounding rock undergoes uneven deformation; Destruction stage: In the late stage of surrounding rock deformation, the rock mass that plays a dominant role in the development of surrounding rock deformation usually has structural fracture characteristics.

[0073] To verify the above laws, taking into account the structural effects existing in the proportional test, and to clearly reflect the dynamic process of the development of transverse and longitudinal cracks and structural fracture during deformation, the width-to-height ratio of the tunnel was enlarged, and a similar simulation test was used to simulate the deformation and failure process of the surrounding rock of the strong mining tunnel.

[0074] In this embodiment, similar simulation tests are conducted based on geometric parameters, rock mass mechanics parameters, and stress field parameters to simulate the deformation and failure processes of the surrounding rock of the strong mining tunnel. The similar simulation tests can be performed using FLAC 3D Numerical simulation software implementation.

[0075] Using FLAC 3DThe roof zoning failure characteristics at different stages obtained by the numerical simulation software are mainly as follows: coordinated deformation characteristics: at the moment of tunnel excavation, the surrounding rock undergoes elastic deformation. At this time, the surrounding rock layers can still maintain synchronous deformation, and mainly synchronous deformation occurs; uncoordinated deformation characteristics: when the deformation exceeds the elastic limit, the shallow surrounding rock undergoes plastic deformation, and the deformation of the shallow surrounding rock layers begins to be unsynchronized; delamination characteristics: as the uncoordinated deformation further develops, transverse cracks will appear between the layers. If the transverse cracks are large in opening, the roof may directly delaminate. Figure 2 (a) is the delamination feature image of the roof partition damage feature, where 1 st Horizontal fractures propagation is the first horizontal crack expansion, and vertical fractures are vertical cracks. The characteristics of the three-hinge arch structure are: the expansion of horizontal cracks will induce the generation of longitudinal cracks. After the cracks are generated, a three-hinge arch will be formed in the roof. During this period, when the hinges in the middle and both ends of the arch are partially crushed, the horizontal force connection between the fractured rock blocks is lost, the roof rock structure is destroyed, and the bearing capacity is completely lost, the roof collapses. Figure 2 (b) is the characteristic image of the three-hinged arch structure with the top plate partition failure characteristics, where 1 st Roof caving is the first time the roof falls. st Formation ofhinged arch structure is the first formation of hinged arch structure.

[0076] At this point, the roof rock layers show coordinated deformation, uncoordinated deformation, delamination and zoning failure characteristics of three-hinge arch structure from shallow to deep. When the longitudinal cracks continue to expand upward, the scope of surrounding rock failure will continue to expand cyclically, forming secondary zoning failure. Figure 2 (c) is the roof partition failure characteristic of secondary partition failure, where 2 nd Horizontal fractures propagation and bed separation refer to the second horizontal crack propagation and interlayer delamination, and vertical fractures refer to vertical fractures. In this embodiment, the coordinated deformation characteristics, uncoordinated deformation characteristics, delamination characteristics, and three-hinge arch structure characteristics are used as the roof partition failure characteristics.

[0077] And through the above-mentioned roof zoning failure characteristics, it can be seen that the formation and upward extension of longitudinal cracks play a dominant role in the formation of roof zoning failure structure. From the perspective of the interaction between the internal structure of the rock stratum, the deformation and failure of the surrounding rock of the strong mining tunnel is a dynamic process in which the mutual squeezing effect of the internal structure of the rock stratum weakens as the longitudinal cracks penetrate, and increases as the rock stratum structure bites each other until the final equilibrium is reached.

[0078] like Figure 3 As shown, Figure 3 The figure shows the time-displacement diagram of the entire deformation and failure process of the surrounding rock in the strong mining tunnel. The instantaneous stage corresponds to the time period 0-t1, with a displacement of 0-S1. The deformation mechanism of the surrounding rock in this stage is the expansion of the rock mass caused by the sudden release of elastic energy and the expansion of transverse cracks caused by the uncoordinated deformation between layers. The displacement includes linear deformation and nonlinear deformation, and the overall deformation is small. The sudden change stage corresponds to the time period t1-t2, with a displacement of S1-S2. The deformation mechanism of the surrounding rock in this stage is the sudden displacement change caused by delamination. This stage lasts for a short time, but the displacement is large. The roof begins to suffer structural damage and the bearing capacity begins to decline. The slow increase stage corresponds to the time period t2-t3, with a displacement of S2-S3. The deformation mechanism of the surrounding rock in this stage is continuous creep deformation. The destruction stage corresponds to the time period t3-t, with a displacement of 0. As the deformation intensifies, the longitudinal cracks move to the deep layer, and the shallow surrounding rock undergoes structural fracture until they interlock to form a three-hinged arch structure and finally stabilize. In this stage, the deformation mechanism of the surrounding rock is the rigid body displacement caused by the fracture of the roof rock layer and its re-stabilization process.

[0079] It should be noted that the principle of the instantaneous stage 0-t1 is the elastic deformation of the surrounding rock at the moment of excavation unloading, but the elastic deformation of rock strata at different levels is inconsistent, so it is called inter-layer inharmonious deformation. However, overall, this deformation is small and has not reached the level of inharmonious deformation, so it is called coordinated deformation.

[0080] In summary, the instantaneous stage corresponds to the coordinated deformation characteristics; the drastic change stage corresponds to the uncoordinated deformation characteristics; the slow increase stage corresponds to the delamination characteristics; and the failure stage corresponds to the three-hinged arch structure characteristics.

[0081] Step S103, determining the key stage of irreversible deformation of the surrounding rock of the strong mining roadway based on the correlation relationship, where the key stage is a stage in the deformation process of the surrounding rock of the strong mining roadway;

[0082] Step S104, determining the critical time for thick layer cross-boundary anchoring based on the key stage;

[0083] Step S104 specifically includes: determining the stage of irreversible deformation of the surrounding rock of the heavily mined roadway based on the roof zoning failure characteristics; using the start time of the irreversible deformation stage as the critical time; wherein the irreversible deformation stage is one of the deformation stages of the surrounding rock of the heavily mined roadway. Furthermore, based on the entire deformation and failure process of the heavily mined roadway, determining that the formation and upward extension of longitudinal cracks play a dominant role in the formation of the roof zoning failure structure, wherein the formation and upward extension of longitudinal cracks occur after the prompt stage.

[0084] In this embodiment, it can be seen from the roof partition destruction structure that the upheaval stage is the key stage for structural damage to occur and the key stage for irreversible deformation of the surrounding rock. The formation and upward extension of longitudinal cracks occur after the instantaneous stage. Therefore, the starting time of the upheaval stage is taken as the critical time, that is, the roof needs to be supported at the critical time at the latest.

[0085] Step S105, determining the support parameters of the anchor rod and the critical anchoring length of the thick layer span, where the critical anchoring length is the sum of the damaged layer of the roof rock mass caused by the mining and excavation process of the surrounding rock working face of the strong mining roadway and the anchoring boundary;

[0086] Among them, determining the support parameters includes: calculating the shear stress on each longitudinal section parallel to the neutral layer when the roof rock beam undergoes pure bending deformation; determining the shear force on each anchor rod based on the shear stress; obtaining the stiffness of the anchor rod and the increment of the anchor rod length; calculating the tension borne by each anchor rod based on the stiffness and the increment of the anchor rod length; and determining the support parameters of the anchor rod based on the shear force and the tension borne.

[0087] In this embodiment, the mechanical analysis method is used to propose the thick layer cross-boundary anchoring principle. The roof rock beam has a tendency to bend and deform under the action of the overburden gravity load. Therefore, according to the principles of geometry, the deformation and damage that may occur when the roof rock beam is compressed and bent are analyzed, and the additional constraints that need to be provided, namely the support parameters, are analyzed.

[0088] Specifically, the following restraints need to be provided during the deformation of the roof rock beam.

[0089] (1) Shear force constraint

[0090] like Figure 4 As shown in the figure, when the roof rock beam undergoes pure bending deformation, shear stress must exist on each longitudinal section parallel to the neutral layer. The anchor rod needs to overcome this shear force to prevent the rock beams from sliding against each other. Figure 4 In the figure, the shear stress at any point k on the cross section at a distance y from the neutral axis when the rock layer is bent, where b is the width of the cross section and h is the height of the cross section. Figure 5 Schematic diagram of the shear force borne by the rock layer when there is no anchor rod. Figure 6 Schematic diagram of the shear force borne by the anchor rod when the rock layer bends when there is an anchor rod.

[0091] According to the shear stress reciprocity theorem, the shear stress on the longitudinal section is equal to the shear stress on the cross section perpendicular to the surface, so we can get .

[0092] In addition, based on the assumption that the roof rock beam has small deformation in pure bending, it can be assumed that the shear stress on the longitudinal section is equal. Therefore, the shear force on a single anchor is , where τ is the shear stress on the longitudinal section, τ' is the shear stress on the cross section perpendicular to the surface, and F s is the cross-sectional shear force; F τ is the anchor shear force, L is the span of the roof rock beam, n is the number of anchors, F is the overburden load, I z is the moment of inertia of the cross section about the neutral axis, h is the height of the cross section, and y is the distance from any point k on the cross section to the neutral axis.

[0093] (2) Tensile force constraint

[0094] like Figure 7 As shown, Figure 7 The diagram of the tension borne by the anchor rod when the rock layer is separated is shown in Figure 2. After the tunnel is excavated, the rock layers tend to move away from each other due to the uncoordinated deformation. In the process of overcoming this trend, the anchor rod is subjected to tensile stress. The tensile force Ft it bears is , where F t is the anchor tension, k is the anchor elastic modulus, and ΔL is the anchor elongation increment.

[0095] Table 1 is a parameter comparison table of steel anchor rods and flexible anchor rods.

[0096] Table 1

[0097]

[0098] As can be seen from Table 1, the flexible anchor rod is superior to the steel anchor rod in many parameters. Therefore, the anchor rod in the technical solution of this application adopts the flexible anchor rod.

[0099] In this embodiment, the support parameters include shear force and tension force as support parameters, and flexible anchor rods are used.

[0100] In this embodiment, the critical anchoring length of the thick layer span of the anchor rod is determined, including: simulating the mining and excavation process of the surrounding rock working face of the strong mining tunnel to obtain the damaged layer and anchoring boundary of the roof rock mass; calculating the sum of the damaged layer and the anchoring boundary; and taking the sum as the critical anchoring length.

[0101] Under strong mining conditions, the surrounding rock is subjected to strong mining pressure caused by tunnel excavation and mining. The surrounding rock is distributed in the order of Excavation Damaged Zone (EDZ), Excavation Influenced Zone (EIZ) and Intact Zone. Among them, EDZ is an annular fracture zone caused by strong mining disturbance caused by tunnel excavation and mining. In this area, the rock structure is severely damaged and the rock mass strength is between the peak strength and the residual strength. Figure 8As shown in the figure, if no external constraints are applied, its bearing capacity will be completely lost; EIZ refers to the surrounding rock with stress lower than the peak strength, less cracks, and undamaged rock structure; Intact zone refers to the surrounding rock in the original rock stress state and is not affected by mining disturbance. Figure 8 In the formula, Peak stength is the peak strength, that is, the maximum stress value reached by the rock mass during the loading process, marking the transition from elastic deformation to nonlinear deformation; Elastic limit is the elastic limit, that is, the end point of the linear segment in the stress-strain curve, and the deformation can be completely recovered after unloading; Plastic boundary is the plastic boundary, that is, the boundary of the region where the stress state in the rock mass reaches the yield condition; Residuale stength is the residual strength, that is, the ability of the rock mass to continue bearing after failure, corresponding to the stable stress value in the softening segment of the stress-strain curve; Strain is the strain, which is a mechanical parameter describing the degree of deformation of the rock mass; Stress is the force per unit area inside the rock mass.

[0102] When the anchor end is arranged in the EDZ, the anchor system cannot provide stable bidirectional constraints to the anchor body, nor is it conducive to the continuous transmission of stress within the anchor body, which easily causes stress concentration. This area is the main control object in the support system; when the anchor end is arranged in the EIZ, the self-bearing performance of the rock mass in this area and deeper can be fully mobilized to achieve deep-shallow linkage. Therefore, the rock mass in this area can serve as the bearing body of the support system.

[0103] In summary, the damaged layer of the roof rock mass caused by working face mining and tunnel excavation and the anchoring boundary are called the critical anchoring length, which will vary with different geological and engineering conditions. The principle of cross-boundary support means that the anchoring range must penetrate the critical anchoring length so that the anchoring end is located in the elastic rock mass deep in the overburden. Figure 9 Schematic diagram of the distribution of EDZ, EIZ, Intact Zone and plastic zone of surrounding rock. Roadway is the roadway, Flexible bolt is the flexible anchor bolt, and principal stress is the principal stress.

[0104] In this way, on the one hand, it can provide the anchor body with stable biaxial stress constraints, so that the surrounding rock is always in a triaxial stress state, enhancing the bearing capacity of the surrounding rock; on the other hand, it is conducive to the continuous transmission of stress and plays a role in fully mobilizing the bearing capacity of the deep surrounding rock.

[0105] In this embodiment, FLAC is used 3D Numerical simulation software analyzes the damaged layer and anchoring boundary of the roof rock mass caused by working face mining and tunnel excavation, thereby obtaining the critical anchoring length.

[0106] The thicker the anchoring layer, the more beneficial it is to the roof rock beam. Specifically, the maximum tensile stress formula for the roof rock beam during pure bending deformation was obtained. The relationship between the maximum compressive stress and thickness was obtained based on the moment of inertia of the cross section about the neutral axis and the maximum tensile stress formula. The relationship is that the maximum tensile stress is inversely proportional to the cube of the thickness, that is, the thicker the anchoring layer, the stronger the bearing capacity of the roof rock beam.

[0107] In this embodiment, considering the pure bending deformation of the roof rock beam, according to the maximum tensile stress theory (first strength theory), the most dangerous point of the roof rock beam is the middle position where the tensile stress causes damage. , and because Therefore, when the external force, structure and constraint conditions of the layered roof rock beam are consistent, the maximum tensile stress is inversely proportional to the cube of the thickness, that is, That is, the thicker the anchoring layer is, the more conducive it is to the bearing capacity of the roof rock beam, where σ is the bending moment, M y is the cross-sectional normal stress.

[0108] Step S106: Support the roof of the heavily mined tunnel based on the critical time, support parameters and thick layer cross-boundary anchoring length.

[0109] In this embodiment, before the early stage of drastic changes in surrounding rock deformation, the roof of the heavily mined tunnel is supported based on support parameters, flexible anchor rods and thick cross-layer anchoring length, thereby reducing the possibility of large deformation of the roof surrounding rock or even sinking of the entire structure.

[0110] The following example illustrates the thick layer cross-boundary support of mining tunnels.

[0111] In this embodiment, Figure 10 The mining of the coal mine shown in FIG. 1 is used as an example to illustrate the mining of the coal mine shown in FIG. - 1 106 = 14 -1 106 goaf, Mined-out 14 -1 104 = 14 -1 104 goaf, Working Face 14 -1 105 = 14 -1 105 working surface, Headentry of 14 -1 106 = 14 -1 106 working face transport tunnel, Headentry of 14 -1 105 = 14 -1The coal pillar (6m in width) in the 105 working face transport tunnel is 6m wide. The mine has an annual production of 1.5 million tons and is currently mining the first mining area of ​​the 4-1 coal seam. The coal seam is buried at a depth of 490m and has a coal thickness of 2.87 to 3.64m. The direct roof is composed of interbedded mud and sand, and the old roof is mainly composed of medium-fine sandstone. The lithologic characteristics of the roof are shown in Table 2. Seven working faces are arranged on the south wing of the first mining area, and the mining is carried out alternately between sections. Due to the failure to properly avoid mine pressure, the goaf-side tunnels have been severely deformed and need to be renovated several times during service. As a result, the mine has 14 -1 The 105 working face will soon face the problem of strong mining in isolated working face mining. -1 105 working face in the same mining area 14 -1 The layout and damage characteristics of the 106 head-end goaf-side tunnel are that the roof surface begins to break, rock blocks fall in the shape of a "net bag", and the shallow surrounding rock undergoes uneven deformation.

[0112] In order to further verify the adaptability of thick layer cross-boundary anchoring technology to the maintenance and control of strong mining tunnels, 14 -1 Industrial test was carried out on the mining conditions of small coal pillar along goaf in isolated island working face of 105 working face. -1 105 lower chute is located at 14 - 1 East side of 106 working face, rectangular section, along 14 -1 104 working face goaf and 4-1 coal floor excavation, net section is 4.6m×3.2m, total length of tunnel is 566m, 14 -1 105 and 14- 1 The coal pillar width of the 104 working face section is 6m, and the mining layout is as follows: Figure 10 shown.

[0113] Determine the support parameters of the anchor rod and the critical anchorage length of the thick layer span. The support parameters are:

[0114] Determining the critical anchorage length: using FLAC 3D Numerical simulation software was used to analyze the damaged layer and anchoring boundary of the roof rock mass caused by working face mining and tunnel excavation. The geometric parameters were 700 meters long, 200 meters wide, and 150 meters high. The upper surface was the stress boundary, and 9.1 MPa was applied to simulate the deadweight of the overlying strata. The lateral stress coefficient was 1.2. The Mohr-Coulomb constitutive model was used, and the minimum size of the tunnel surrounding rock unit cell was 0.2×0.2×0.2m. The small coal pillar width was 6m. Assuming that 14 -1 104 and 14 -1 106 working face is mined backward, and 14 -1 Tunnel excavation and retreat mining are carried out at the 105 working face.

[0115] Table 2 is the lithologic characteristics of the roof

[0116] Table 2

[0117]

[0118] Among them, FLAC 3D The simulation process of the numerical simulation software specifically includes: (1) implementing and initializing the in-situ stress field; (2) mining the working face 14 at intervals of 20 meters in the mining direction in each cycle. -1 104, until the entire working face is completely mined; (3) Use the same procedure to mine the working face 14 -1 106; (4) Excavate along the roadway inclination and apply different support schemes, including no support scheme, original support scheme and thickened cross-boundary support scheme; (5) Perform solution steps and calibrate deformation characteristics as needed; (6) Mining working face 14 -1 105.

[0119] The following explanations are given of the original support scheme using left-handed unreinforced threaded steel anchor rods in combination with anchor cable support and the thick layer cross-boundary anchor support scheme using flexible anchor rod support.

[0120] (1) Left-handed non-longitudinal reinforced threaded steel anchor rod combined with anchor cable support

[0121] As shown in Figure 11 (a) and (b), cable bolts and rebar bolts are left-handed unreinforced threaded steel anchor rods. The original support scheme uses left-handed unreinforced threaded steel anchor rods combined with cable anchor support. Six Φ20×2200mm threaded steel anchor rods are arranged in each row of the tunnel roof, with a row spacing of 800×1000mm; two or three Φ18.96×6500mm anchor cables are arranged every two rows of anchor rods, forming a "3-2" arrangement; four Φ20×2200mm threaded steel anchor rods are arranged in each row of the side, with a row spacing of 1000mm and different spacing. The preload torque of the threaded steel anchor rods is not less than 60kN, and the preload force of the anchor cables is not less than 300kN.

[0122] (2) Thick layer cross-border anchor support scheme adopts flexible anchor support

[0123] As shown in Figure 12 (a) and (b), flexible bolts are used for thick-layer cross-boundary anchorage support. Five Φ21.8×4300mm flexible anchors are arranged in each row of the top plate, with a row spacing of 950mm×1000mm. Two or three Φ21.8×6300mm anchor cables are arranged every two rows of flexible anchors, with different spacing and a row spacing of 2000mm. The preload force of the flexible anchors is not less than 200kN, and the preload force of the anchor cables is not less than 300kN.

[0124] At the moment of tunnel excavation, i.e., the instantaneous stage, the above-mentioned left-handed unreinforced threaded steel anchor rods combined with cable anchor support and thick-layer cross-boundary anchor support scheme were simulated using flexible anchor rod support.

[0125] The simulation results of the distribution characteristics of the surrounding rock stress field and displacement field under different supports are analyzed below.

[0126] In order to clearly reflect the influence of thick-layer cross-boundary anchoring technology on the surrounding rock support effect of strong mining tunnel, without considering the effect of original rock stress, the stress field and displacement field distribution characteristics of the three types of support in the surrounding rock are analyzed, such as Figure 13 As shown in the figure, without considering the original rock stress, the stress field of the anchor support mainly shows the following characteristics:

[0127] Figure 13 (a) is the stress field and displacement field without support, where the tensile stress zone is the tensile stress zone. Figure 13 (a) It can be seen that the roof surrounding rock is directly exposed to the tensile stress environment, and the maximum tensile stress value reaches 0.02MPa.

[0128] Figure 13 (b) is the support stress field and displacement field under the original support condition, where the Compressive stress zone is the compressive stress zone. Figure 13 (c) shows the support stress field and displacement field under thick cross-boundary anchor support, where stress driven effect is the stress driven effect. Figure 13 (b) and Figure 13 (c) It can be seen that tensile stress concentration occurs in the deep range of the anchoring end of the left-handed threaded steel anchor rod and the flexible anchor rod, with the maximum tensile stress values ​​being 0.11MPa and 0.36MPa respectively. The tensile stress values ​​gradually decrease as they go deeper into the tunnel roof and away from the anchoring end of the anchor rod; obvious compressive stress concentration areas appear in the range from the anchoring end to the tail end of the left-handed threaded steel anchor rod and the flexible anchor rod, with the maximum compressive stress values ​​being 0.09MPa and 0.34Mpa respectively.

[0129] Under the schemes of no support, original support and thick cross-border anchoring support, the roof subsidence after the tunneling was stabilized was 127.4mm, 58.9mm and 50.8mm respectively. Compared with the schemes of no support and original support, the roof subsidence was reduced by 60.13% and 13.75% respectively when thick cross-border anchoring support was used, which can effectively suppress the roof subsidence.

[0130] In summary, for the original support and thick cross-border anchor support, the maximum tensile stress value is close to the maximum compressive stress value, and the compressive stress area and the tensile stress area are approximately symmetrically distributed along the end of the anchor rod. Compared with the surrounding rock stress field distribution when there is no support, it can be seen that the optimization effect of anchor support on the surrounding rock stress field mainly includes two aspects: tensile stress driving effect and compressive stress transfer effect. Moreover, with the increase of anchor rod length, the tensile stress driving effect and compressive stress continuous transfer effect become more obvious.

[0131] Since sedimentary rock layers are brittle materials, the coordinated deformation coefficient is used as an indicator to determine whether the rock layer has undergone brittle fracture. When obvious uncoordinated deformation occurs, it is determined that the layered roof has undergone brittle fracture at the maximum curvature radius; otherwise, it is considered to be in an elastic deformation state. In addition, when the curvature distribution of the movement curves of adjacent rock layers is uneven, it is determined that separation has occurred between adjacent rock layers. Based on the above judgment, the displacement data of the roof rock layer in the range of 0 to 10.8m is extracted (extracted every 0.2m), and the partitioned fracture characteristic range of the roof layer structure is drawn, such as Figure 13 shown.

[0132] Figure 14 (a) is the rock movement curve of the unsupported roof, where Distance from the left rib is the distance from the left side of the roadway, Deformation is the deformation, Bed separation is the stress change between the top and bottom plates of the rock layer (or coal seam) and the main layer, Bed breaking is the bedrock fracture, 1 is the coordinated deformation zone, 2 is the separation zone, and 3 is the three-hinge arch structure zone. According to the different degrees of bending and sinking of the layered roof, the rock layer structure of the unsupported roof can be divided into the coordinated deformation zone, the separation zone and the three-hinge arch structure zone. ① In the range of 0 to 3.2 m, the layered roof undergoes uncoordinated deformation. Since sedimentary rock is brittle, it is believed that the rock strata in this area undergo brittle fracture along the middle, forming a three-hinge arch structure. The rock strata in this area are broken by horizontal compression, and the broken rock blocks bite each other, forming a three-hinge arch structure as a whole. Considering that there is no external support constraint, the bearing capacity of the rock strata in this area is basically lost. ② In the range of 3.2 to 9.4 m, the basic shape of the layered roof can still be maintained, but the sinking deformation in the middle of the rock strata is more prominent, and there is delamination in some areas. ③ In the range of 9.4 to 10.8 m, the layered roof undergoes coordinated deformation. The rock strata in this area are under pressure as a whole. The layered structure of the roof is intact and has a certain bearing capacity, but it tilts and sinks as a whole toward the goaf.

[0133] like Figure 14(b) shows roof caving, which refers to roof collapse. The zonal failure characteristics of the roof strata under the original support scheme differ significantly from those under no support in terms of zonal failure range, rock fracture morphology, and roof displacement. ① Within the 0-3.8m range of the roof, the layered roof undergoes uncoordinated deformation, forming a three-hinged arch structure overall, but its degree of sinking and tilting is significantly less than that under no support. ② Within the 3.8-6.6m range of the roof, the layered roof undergoes bending deformation. From the goaf side to the solid coal side, the tendency of the layered roof to squeeze each other becomes increasingly obvious, and there is a clear tendency for the roof to separate near the goaf. ③ Within the 6.6-10.8m range of the roof, the layered roof undergoes coordinated deformation, the rock strata are relatively intact, and the rock strata tilt is less than that under no support.

[0134] like Figure 14 As shown in (c), under the thick cross-border anchor support, ① within the range of 0 to 3.6 m of the roof, the rock stratum undergoes uncoordinated deformation, and the overall structure is in a three-hinged arch state; ② within the range of 3.6 to 6.6 m, the rock stratum undergoes bending deformation, and the degree of mutual squeezing of the roof rock strata in the middle of the tunnel is relatively weakened, with a tendency of separation; ③ within the range of 6.6 to 10.8 m, the rock stratum mainly undergoes coordinated deformation, the rock stratum bearing structure maintains its intact form, and the rock stratum inclination is less than that in the original support and no support cases.

[0135] In summary, from the perspective of roof structural characteristics, both the thick cross-border anchor support and the original support have three-hinged arch structure areas, separation areas, and coordinated deformation areas, but their structural characteristics show significant differences. The main difference is that under the original support scheme, the mutual squeezing trend of the layered roof structure from the goaf side to the solid coal side becomes more obvious, and the maximum subsidence occurs in the roof on the goaf side. Under the thick cross-border anchor support scheme, the mutual squeezing trend of the layered roof structure from the goaf side to the solid coal side first increases and then decreases, with the maximum subsidence occurring in the middle of the roof. The roof near the goaf side is significantly squeezed, and the subsidence trend is significantly weakened. The goaf side roof is significantly controlled by the thick cross-border technology. Combined with the rock stratum displacement characteristics, the maximum roof subsidence under the thick cross-border anchor support and the original support scheme is 50.8mm and 58.9mm, respectively. It can be seen that the more obvious the rock stratum mutual squeezing trend, the better the structural integrity, and the more conducive to deformation control.

[0136] When the anchoring length reaches 4m, the stress field of the surrounding rock supported by the anchor rod changes significantly, and the tensile stress transfers from the shallow surrounding rock to the deep overburden. At the same time, the effective compressive stress zones of the roof are superimposed on each other, and the compressive stress control range expands upward. In addition, in terms of roof structural characteristics, the sinking trend of the roof on the goaf side is significantly weakened, and the thick layer cross-border technology has a significant control effect on the roof structure, and the structural integrity is good.

[0137] In summary, compared with no support and the use of left-handed unreinforced threaded steel anchor support, the use of flexible anchor support can improve the stress field of the support structure and optimize the structural characteristics of the roof to a certain extent.

[0138] In order to further verify the thick layer cross-boundary anchoring technology, the thick layer cross-boundary anchoring scheme shown in Figure 11 (b) was used to headentry of 14 -1 105 is supported for further field application.

[0139] At 14 -1 Five measuring points are arranged in the lower drift of the 105 working face to carry out on-site monitoring of the tunnel surface displacement, surrounding rock separation, anchor (cable) stress, and roof rock crack development information.

[0140] 14 -1 105 After the headentry is opened for 200m, a roof displacement observation station is arranged every 10m. The final roof subsidence at the five observation stations is 28mm, 15mm, 67mm, 80mm and 50mm respectively. Figures 15(a) and 15(b) show the roof displacement at 14 -1 The roof deformation of the air inlet chute at the 105 working face was measured, where "Deformation" refers to deformation, "Distance to the excavation face" refers to the distance from the excavation face, "Deformation Rate" refers to the deformation rate, "P1", "P2", "P3", "P4", and "P5" refer to measuring points, "MonitoringDeformation" refers to rock and soil deformation monitoring, and "Average deformation" refers to the average deformation. Based on the different deformation rates, the surrounding rock deformation exhibits obvious stage characteristics, and the roof subsidence can be divided into a period of intense deformation, a period of slowing deformation, and a period of stable deformation.

[0141] Stage I: Severe Deformation: Within six days of tunnel excavation, the rock mass undergoes severe deformation, with roof subsidence rates ranging from 5 to 10 mm / day, peaking at 13 mm / day. This roof subsidence is primarily due to the release of elastic strain energy after displacement constraints are removed, and rock mass deformation is primarily elastic.

[0142] Stage II: Deformation Stabilization: Within 6 to 15 days of tunnel excavation, rock mass deformation slows, with the roof sinking to 3 to 5 mm / day. Deformation during this stage is primarily due to the tensioning and further expansion of internal cracks in the roof, leading to the re-stabilization of the rock structure. As deformation grows, the rock mass enters a plastic state, and microcracks increase in size. Rock mass deformation is primarily due to irreversible deformation, such as plastic deformation and rigid body displacement.

[0143] Stage III: Deformation stabilization stage. 15 days after tunnel excavation, stress redistribution is completed, deformation tends to be stable, and the sinking rate drops to 0-2 mm / d.

[0144] Figure 15(c) shows the three stages of roof subsidence, where "Monitoring station" represents the monitoring station, "Proportion of deformation" represents the deformation ratio, and "Total deformation" represents the total deformation. As shown in Figure 15(c), the greater the total roof deformation, the greater the proportion of periods of intense deformation and periods of slowing deformation. After tunnel excavation, the preload of the flexible anchors failed to effectively control the bending deformation of the shallow surrounding rock, leading to delamination of the roof surrounding rock and further crack development. This prevented the roof surrounding rock from forming an effective load-bearing structure, leading to further roof deformation. Therefore, the optimal period for surrounding rock control is during the initial stage of tunnel excavation. Applying high preload and thick anchoring at this time is an effective means of controlling the roof surrounding rock.

[0145] A borehole peep instrument was used to observe the damage characteristics of the roof partitions. Figure 16 This is the internal rock structure of the roof of borehole 1. The roof only shows extrusion-type fractures within the 0-0.2m range at the entrance, and longitudinal cracks are distributed between 2.1 and 2.2m. No significant cracks or weak planes are observed in the remaining layers, and the overall state is intact. The internal rock structure and failure pattern of the roof of borehole 2 are similar to those of borehole 1. Aside from surrounding rock fractures within the 0-0.3m range at the entrance and longitudinal cracks between 2.1 and 2.2m, the deep rock mass remains largely intact, with no significant circumferential cracks observed, and no delamination of the roof has been observed.

[0146] In summary, after implementing thick-layer cross-boundary anchoring support, longitudinal crack development was limited to the roof depth of 2.3 m or less. Except for some extrusion-type fractures at the opening, the remaining layers were primarily naturally developed coal-rock interfaces and weak planes. The surrounding rock was relatively intact, with no obvious zonal failure characteristics.

[0147] It can be seen that the improvement effect of thick cross-boundary anchoring on the stress field and displacement field of surrounding rock support is mainly reflected in the following three aspects:

[0148] (1) Tensile stress driving effect. As shown in Figure 17 (a), where Distance from the roof surface is the distance from the roof surface, Value of Tensile Stress is the tensile stress, and Crifical Anchorage Length is the critical anchorage length, when the anchor end is arranged deeper than the critical anchorage length, the tensile stress in the shallow rock mass will be transferred to the deep overburden. At this time, the tensile stress in the shallow range of the critical anchorage length is no support, original support, and optimized support, while the tensile stress in the deep range of the critical anchorage length is optimized support, original support, and no support. Considering that the deformation resistance and bearing capacity of the surrounding rock gradually increase with the depth of the surrounding rock, the shallow rock mass at the critical anchorage length is sensitive to the tensile stress, while the surrounding rock at the deep range is stable and has strong bearing capacity, and is not very sensitive to the tensile stress. In this way, the application of cross-border support can play a dual role of stress equalization and stress dilution, avoiding the direct tensile stress on the shallow damaged rock mass and effectively controlling the deformation of the shallow surrounding rock, where 1 is the tensile stress value when thick-layer cross-border anchoring support is adopted; 2 is the tensile stress value when combined support with anchor rods and cables is adopted; 3 is the tensile stress value when there is no support; 4 is the strain curve when thick-layer cross-border anchoring support is adopted; 5 is the strain curve when combined support with anchor rods and cables is adopted; 6 is the strain curve when there is no support.

[0149] (2) Continuous transmission of compressive stress. As shown in Figure 17 (b), where Compressive stress Control boundary is compressive stress, after applying cross-boundary anchoring, the effective compressive stress zones of the top plate are superimposed on each other, and the compressive stress control range expands upward, basically covering all rock masses shallower than the critical anchoring length. In this way, on the one hand, it can provide bidirectional constraints to the rock mass between the anchor end and the anchor tail, so that it is in a three-dimensional compressive state, strengthen the bearing capacity of the surrounding rock, and prevent further expansion of rock damage; on the other hand, it is conducive to the bidirectional continuous transmission of stress in the anchor body, so that it has a direct connection with the deep overburden, and mobilizes a larger range of rock masses to participate in the bearing, realizes stress homogenization and deep-shallow linkage of displacement, and further limits the displacement of shallow rock masses. Among them, 7 is the compressive stress value when thick cross-boundary anchoring support is used; 8 is the compressive stress value when anchor cable combined support is used; 9 is the compressive stress value when there is no support; 10 is the rock mass deformation resistance.

[0150] (3) Strain zero point downward shift feature. The displacement of the roof surface is equal to the accumulation of linear strain from the strain zero point deep in the surrounding rock to the surrounding rock surface. Under the action of thick cross-boundary anchoring, the tensile stress is transferred to the depth of the surrounding rock. This causes the strain zero point to gradually shift downward in the shallow range of the critical anchoring length, reducing the deformation range of the surrounding rock, thereby achieving effective control of the surrounding rock deformation.

[0151] In summary, compared with the original support scheme, the optimization and improvement of the roof structural characteristics by the thick-layer cross-border anchoring scheme are mainly reflected in the following three aspects: First, the optimization of the roof zoning structure range. Under the thick-layer cross-border anchoring scheme, the range of the roof three-hinged arch structure area is 3.6m, which is 0.2m less than that of the original support scheme, which is more conducive to the load-bearing and stress optimization of the roof structure. Second, the difference in structural morphology. Under the thick-layer cross-border anchoring scheme, the roof near the goaf is significantly squeezed and the sinking trend is significantly weakened. The goaf-side roof is significantly controlled by the thick-layer cross-border anchoring technology, and the structural integrity is better. Third, the structural stress state is different. Under the thick-layer cross-border anchoring scheme, the rock layer movement curve is more dense, indicating that the rock layers are significantly squeezed by each other, the three-way compression effect is significant, and the ultimate bearing capacity of the rock layer is further improved. Therefore, the control effect of the thick-layer cross-border anchoring technology on roof deformation is mainly reflected in the above three aspects: first, the optimization of the roof zoning fracture structure, second, the optimization of the structural morphology, and third, the optimization of the rock layer structural bearing capacity.

[0152] The terms "comprises," "comprising," or any other variations thereof are intended to cover non-exclusive inclusion such that a process, method, article, or apparatus that comprises a list of elements includes not only those elements but also other elements not expressly listed or inherent to such process, method, article, or apparatus.

[0153] The above description is merely a preferred embodiment of the present application and an illustration of the technical principles employed. Those skilled in the art should understand that the scope of application involved in this application is not limited to the technical solutions formed by a specific combination of the above-mentioned technical features, but should also cover other technical solutions formed by any combination of the above-mentioned technical features or their equivalents without departing from the aforementioned application concept. For example, a technical solution formed by replacing the above-mentioned features with (but not limited to) technical features with similar functions applied for in this application.

Claims

1. A method for cross-boundary support of thick layers in a strong mining tunnel, characterized by: include: Obtain geometric parameters, rock mass mechanical parameters and stress field parameters of strong mining tunnels; The deformation and failure processes of the surrounding rock of the strong mining tunnel are simulated based on the geometric parameters, rock mass mechanics parameters and stress field parameters, and the correlation between each deformation stage of the surrounding rock of the strong mining tunnel and the failure characteristics of the roof partition is obtained; Determine a critical stage in which irreversible deformation of the surrounding rock of the strongly mining roadway occurs based on the correlation relationship, wherein the critical stage is a stage in the deformation process of the surrounding rock of the strongly mining roadway; Determining the critical time for thick layer cross-boundary anchoring based on the key stages; Determining the support parameters of the anchor rod and the critical anchoring length of the thick layer span, wherein the critical anchoring length is the sum of the damaged layer of the roof rock mass caused by the mining and excavation process of the surrounding rock working face of the strong mining roadway and the anchoring boundary; The roof of the heavily mined tunnel is supported based on the critical time, support parameters and thick layer cross-boundary anchoring length.

2. According to the method according to claim 1, the deformation stages of the surrounding rock of the strong mining tunnel include instantaneous stage, sudden change stage, slow increase stage and destruction stage.

3. The method according to claim 2, characterized in that The roof zoning destruction characteristics of the surrounding rock of the strong mining tunnel are as follows: the roof rock layers show coordinated deformation characteristics, uncoordinated deformation characteristics, delamination characteristics and three-hinge arch structure characteristics from shallow to deep.

4. The method according to claim 3, characterized in that The relationship between each stage and the roof partition failure characteristics is as follows: The instantaneous stage corresponds to the coordinated deformation characteristics; the drastic change stage corresponds to the uncoordinated deformation characteristics; the slow increase stage corresponds to the delamination characteristics; and the destruction stage corresponds to the three-hinged arch structure characteristics.

5. The method according to claim 1, characterized in that The critical time for thick layer cross-boundary anchoring is determined based on the key stages, including: Determining the stage where the surrounding rock of the intense mining tunnel produces irreversible deformation based on the roof zoning damage characteristics; The starting time of the irreversible deformation stage is taken as the critical time; The irreversible deformation stage is one of the deformation stages of the surrounding rock of the intense mining tunnel.

6. The method according to claim 1, characterized in that Determine the anchor support parameters, including: Calculate the shear stress on each longitudinal section parallel to the neutral layer when the roof rock beam undergoes pure bending deformation; determining a shear force applied to each anchor rod based on the shear stress; Obtain the stiffness of the anchor rod and the increment of the anchor rod length; Calculating the tension borne by each anchor rod based on the stiffness and the increment of the anchor rod length; The support parameters of the anchor rod are determined based on the shear force and the tensile force.

7. The method according to claim 1, characterized in that Determine the critical anchorage length of the anchor bolt across thick layers, including: The mining and excavation process of the surrounding rock working face of the strong mining tunnel was simulated to obtain the damaged layer and anchoring boundary of the roof rock mass; Calculating the sum of the damaged layer and the anchoring boundary; The sum is taken as the critical anchoring length.

8. The method according to claim 6, characterized in that The anchor rod is a flexible anchor rod.

9. The method according to claim 7, characterized in that Also includes: Obtain the formula for the maximum tensile stress when the roof rock beam undergoes pure bending deformation; The relationship between the maximum compressive stress and thickness is obtained based on the moment of inertia of the cross section about the neutral axis and the maximum tensile stress formula; The relationship is that the maximum tensile stress is inversely proportional to the cube of the thickness, that is, the thicker the anchoring layer, the stronger the bearing capacity of the roof rock beam.

10. The method according to claim 1, characterized in that Based on the whole process of deformation and failure of the strong mining tunnel, it is determined that the formation and upward extension of longitudinal cracks play a dominant role in the formation of the roof partition failure structure, among which the formation and upward extension of longitudinal cracks occur after the instantaneous stage.

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