Method for determining limit value of roadway roof separation of weakly cemented mining stratum

Abstract

The invention discloses a weakly cemented mining stratum roadway roof separation limit value determination method, relates to the technical field of mine roadway damage surrounding rock anchor net cable support and roof stability monitoring, and focuses on roadway surrounding rock roof support and monitoring of a weakly cemented coal mining stratum. The most typical damage form is sudden roadway roof fall and roadway integral convergence, and the life safety of mine engineering technicians is greatly threatened. The roof separation value is the relative displacement between two measuring points in the surrounding rock and between the surrounding rock at the end of the anchor rod and the surface of the roadway roof, belongs to the mine pressure display of the mine roadway, and is an important index for evaluating the anchor net cable supporting effect. The anchor net and anchor cable roof separation value based on the excavation disturbance rock damage degree is accurately calculated and monitored, and the method has important significance for guaranteeing mine safety production.

Description

TECHNICAL FIELD

[0001] The present application relates to the technical field of mine roadway damage surrounding rock anchor net cable support and roof stability monitoring, in particular to a weakly cemented mining stratum roadway roof separation limit value determination method. BACKGROUND

[0002] The weakly cemented coal stratum rock mass has low mechanical strength and is easy to disintegrate. The roof and floor and two sides of the coal mine roadway surrounding rock have complex and various failure mechanisms, and have great safety hazards. The support of the coal mine roadway is mainly anchor rod and anchor cable support. Due to the comprehensive influence of high ground stress, mining stress and pressure water pressure on the underground mine engineering, the design specification and effect evaluation system of the anchor rod and anchor net cable support reinforcement treatment under the existing technical conditions are not perfect, and need to be verified and monitored by engineering practice.

[0003] At present, the determination of the mine roadway roof separation value is mainly realized by installing a roof separation indicator. The sinking displacement of the anchor cable section and the anchor rod section is reflected based on the display value of the inner and outer measuring cylinders of the roof separation indicator, and the roof separation of the anchor rod / anchor cable support range is monitored. However, the sinking displacement value obtained by this measurement method does not consider and cannot monitor the deformation of the roadway roof stratum outside the anchoring range, which belongs to the generalized range of the roof separation value. Since the coal mining site must prevent roof fall accidents, the critical value of the roof separation cannot be obtained by actual measurement.

[0004] Further, the theoretical calculation method for the critical value of the roadway roof separation is: the critical value of the anchor rod anchoring section + the critical value of the anchor cable anchoring section + the critical value outside the anchoring section, wherein the critical value outside the anchoring section is simplified as the deflection deformation value of the anchor cable anchoring rock beam (simply supported beam). Although this method considers the deformation of the roadway roof stratum outside the anchoring range, only the bending moment deformation caused by the vertical uniform load is considered in the process of simplifying the simply supported beam, and the deflection caused by the ground stress is not fully considered, so the obtained separation critical value has limitations.

[0005] Further, for the roof subsidence leading to roof fall accidents, the use of numerical simulation to determine the critical value of separation usually takes the plastic zone range of shear failure as the basis. This method only considers the influence of disturbance pressure load, and less considers the tensile and compressive instability deformation of strength and stiffness parameters, so the limit value of the roadway roof deformation separation calculated according to the plastic zone is not accurate.

[0006] The existing technical method is mostly measured in the case of good support state, which is different from the actual coal mining disturbance roadway surrounding rock instability deformation, and the obtained roof separation value is theoretically much smaller than the engineering actual critical value. SUMMARY

[0007] In order to solve the above problems, the weakly cemented mining stratum roadway roof separation limit value determination method can improve the accuracy of roadway roof separation limit value calculation and monitoring research results.

[0008] To achieve the above object, the weakly cemented mining stratum roadway roof separation limit value determination method comprises the following steps:

[0009] Step S1: obtaining the mechanical parameters of the weakly cemented stratum roadway surrounding rock block by indoor test, the mechanical parameters including density, elastic modulus, Poisson's ratio, internal friction angle, cohesion, compressive strength and tensile strength, the indoor mechanical parameter determination experiment including uniaxial compression, Brazilian splitting and variable angle shear test, and the determination of compressive strength, tensile strength and shear strength is completed;

[0010] Step S2: establishing a small model of the same size as the indoor experiment in step S1 by using the UDEC Trigon model, and inversely calculating the reasonable contact surface parameter value and reasonable block parameter value of the small model established by using the UDEC Trigon model according to the stress-strain curve of the uniaxial compression test and the stress-strain curve of the Brazilian splitting test in step S1;

[0011] Step S3: substituting the rock mass mechanical parameters obtained by the indoor test in step S1 into the empirical formula of the elastic theory and the strength theory to determine the rock mass mechanical parameters;

[0012] Step S4: establishing a uniaxial compression large model with a width of 1m and a height of 2m and a Brazilian splitting large model with a diameter of 1m by using the UDEC Trigon model, substituting the reasonable contact surface parameter value and reasonable block parameter Trigon model parameter determined in step S2 into the large model, and inversely calculating the rock mass contact surface parameter value and block parameter value by using the rock mass mechanical parameters determined in step S2 for the contact parameters and block parameters of the uniaxial compression large model and the Brazilian splitting large model;

[0013] Step S5: verifying the rationality of the rock mass contact surface parameter value and block parameter value, establishing a numerical model of the stability of the surrounding rock of the roadway by using the UDEC Trigon module, continuously iterating and adjusting the rock mass contact surface parameter value and block parameter value determined in step S4 in the numerical model of the stability of the surrounding rock of the roadway until the damage degree of the simulated surrounding rock of the roadway is consistent with the engineering site, and determining the reasonable block parameter and contact surface parameter of the model rock mass;

[0014] Step S6: establishing a weakly cemented rock mass damage degree characterization method, inputting the rock mass block parameter and contact surface parameter of the roadway surrounding rock model after iteration and trial correction in step S5 in the UDEC Trigon model, and inversely calculating the damage-strain relationship curve of the rock mass sample under triaxial compression state;

[0015] Step S7: Study the instability mechanism of the roof rock mass of the weakly cemented surrounding rock roadway;

[0016] Step S8: Simulate the critical damage value of the anchor bearing body of the roof of the weakly cemented surrounding rock roadway under the influence of mining;

[0017] Step S9: Propose a damage discrimination method for the instability of the anchor bearing body of the roof of the weakly cemented surrounding rock roadway;

[0018] Step S10: To ensure the accuracy of the method for calculating the critical value of the roof separation of the roadway, a support structure breakage discrimination method is proposed to characterize the stability of the anchor bearing body of the roof of the roadway by the stress characteristics of the support structure. When the support structure breaks, it means that the subsidence of the anchored rock reaches the limit. The support structure includes anchor rods or anchor cables;

[0019] Step S11: Optimize the anchor rod or anchor cable structure unit;

[0020] Step S12: Establish a UDEC Trigon model of the mining roadway under actual engineering conditions, input the boundary conditions of the calculation model, and use the corrected mechanical parameters of the rock mass as the mechanical parameters of the rock mass in the model. The block uses an elastic model, and the contact surface uses a Mohr-Coulomb model;

[0021] Step S13: Establish a roof separation simulation scheme and a separation limit value criterion: establish a model and assign parameters → apply stress boundary conditions according to different mining conditions → calculate the original rock stress → excavate the roadway to stability;

[0022] Step S14: According to the UDEC Trigon model, simulate the surrounding rock response law of the anchor rod and anchor cable supported mining roadway under the influence of mining, respectively assign different support pressure coefficients K, the larger the K value represents the greater the pressure applied, and simulate the stress characteristics of the surrounding fissure field, displacement field and anchor rod or anchor cable of the anchor net cable supported roadway;

[0023] Step S15: Arrange measuring lines on both sides, roof and floor of the UDEC simulated roadway model, and extract the displacement data to obtain the surface displacement and deep displacement of the roof and floor of the roadway;

[0024] Step S16: Based on the UDEC Trigon model, simulate the surrounding rock response law of the anchor rod and anchor cable supported mining roadway under the influence of mining, obtain the deformation characteristics of the surrounding rock of the working face mining roadway, and obtain the separation limit value of the anchor rod and anchor cable anchoring section at the time of instability from the instability conditions of the roadway. The separation limit value of the roof of the roadway is the sum of the separation limit values of the anchor cable and anchor rod anchoring section. Take a reasonable safety factor to obtain the corresponding separation limit values of the anchor rod anchoring section and the anchor cable anchoring section

[0025] Preferably, in step S1, the uniaxial compression test uses a cylindrical specimen with a width of 50mm and a height of 100mm, the Brazilian splitting test and the variable-angle shear test use a cylindrical specimen with a diameter of 50mm.

[0026] Preferably, in step S2, the numerical model shear modulus G and bulk modulus K are obtained according to the relationship between the elastic modulus E and the Poisson's ratio v in S1:

[0027]

[0028] The contact surface method phase stiffness and shear stiffness in the numerical model are obtained from the following equations:

[0029]

[0030] In the formula, ΔZ min is the minimum side length of the block, k n is the contact surface method phase stiffness, k s is the contact surface shear stiffness.

[0031] The method of numerically inverting the mechanical parameters of the UDEC Trigon model judges the inversion effect by whether the change trend of the axial stress-strain curve of the rock obtained by the laboratory test of the rock block specimen in step S1 and the model simulation is similar;

[0032] The contact surface uses the Residual yield criterion, and the block uses an elastic model to perform a scale numerical simulation experiment on the uniaxial compression and Brazilian splitting standard specimens, and the trial-and-error method is used to continuously adjust the parameters. The numerical calculation results are compared with the laboratory test results through the stress-strain curve of the rock simulation test. When the curve trend is basically consistent, the micro-mechanical parameters of the rock block inverted by the UDEC Trigon model are accurate.

[0033] Preferably, in step S3, the rock mass elastic modulus is obtained by the following formula:

[0034]

[0035] Wherein, E m is the rock mass elastic modulus, E r is the rock block elastic modulus, and RQD is the rock quality index. The uniaxial strength of the rock mass is obtained by the following formula:

[0036]

[0037] Wherein, σ1 is the maximum principal stress at the time of rock mass failure, σ3 is the minimum principal stress at the time of rock mass failure, σ ci is the uniaxial strength of the rock block, m b , and s is the Hoek constant.

[0038] Preferably, in step S6, the Trigon model damage variable Di is defined as the ratio of the length of the failed block contact to the length of all blocks:

[0039]

[0040] wherein L s is the cumulative tensile fracture length at different stages, L n is the cumulative shear fracture length at different stages; and L is the cumulative fracture length.

[0041] Preferably, in step S7, the evolution process is divided into four stages, specifically including an initial damage stage in which the damage variable value is small and can be ignored, a damage stable development stage when the stress value reaches 20% to 40% of the peak stress, a damage rapid development stage when the stress value increases to 70% to 90% of the peak stress, and a damage destruction stage in which the damage variable value tends to be stable; when the stress value is 80% to 90% of the peak stress, the damage variable increases slowly and macroscopic cracks appear, which is the instability stage of the rock mass.

[0042] Preferably, in step S8, the confining pressure in the UDEC Trigon damage evolution model is gradually increased to the confining pressure of the actual roadway roof mining, the critical damage variable of the triaxial compression test when the simulated specimen is destroyed is calculated, and the stress value is near 70% to 80% of the peak stress.

[0043] Preferably, in step S8, the damage variable changes under different confining pressures are fitted to establish a critical damage instability equation Di 临 = f(σ), and the rock mass fracture under different confining pressures is monitored in real time by the critical instability equation in the UDEC Trigon damage evolution model, and the damage variable Di value at this time is calculated in real time; when the real-time calculated damage variable Di value exceeds the critical damage variable in the rock mass instability stage in step S7, instability is determined.

[0044] Preferably, in step S11, optimizing the anchor rod or anchor cable structure unit specifically includes the following steps:

[0045] S111: On the basis of the existing Cable unit in UDEC, the anchor cable breaking failure criterion U(i) ≥ Umax(i) is added, wherein U(i) represents the length of any unit of the anchor cable, which is detected by the existing module of UDEC, and Umax(i) is the limit length of the corresponding unit, which is artificially set;

[0046] S112: The Rockbolt structure unit is optimized, and the Rockbolt failure breaking criterion ε pl ≥ tfstrain; in the formula, tfstrain is artificially set; and ε plThe total tensile yield strain of any element in the anchor bolt is obtained from existing modules of UDEC:

[0047]

[0048] In the formula, θ is the axial strain, d is the anchor diameter, and θ is the axial strain. pl Let L be the average rotation angle of the component, and L be the length of the component.

[0049] If the total tensile yield strain ε of any element of the anchor bolt is... pl When the tensile failure strain limit tfstrain is exceeded, the stress and bending moment of the anchor bolt component will drop to 0 instantaneously, indicating that the anchor bolt has broken at that point. In the model, the anchor cable is simulated using Cable elements, and the anchor bolt is simulated using Rockbolt elements.

[0050] Preferably, in step S16, the delamination value of the top plate is output by simulating top plate delamination under different stress concentration conditions, both inside and outside the anchorage section. When the elongation of the anchor rod exceeds the elongation at break, it indicates that the anchorage material has failed or the damage variable Di value of the surrounding rock within the anchorage range of the support structure exceeds the critical damage value Di. 临 At this point, the delamination value of the top slab is the limit value.

[0051] Therefore, the present invention employs the above-mentioned method for determining the roof delamination limit value in weakly cemented mining strata, which has the following beneficial effects:

[0052] (1) The instability of the anchoring bearing body proposed in this invention includes the surrounding rock and the anchor rods or anchor cables supporting the surrounding rock. This is different from the shortcomings of the existing technology that only considers the single aspect of rock mass instability. It considers the influence of the development of internal micro-rock mass fissures and the fracture failure of external macro-anchoring materials on the instability of the surrounding rock of the roadway.

[0053] (2) In simulating the instability of surrounding rock tunnels, this invention differs from the existing technology in that it directly inputs the mechanical parameters of the rock block. It considers that the mechanical parameters of the rock mass can be obtained by further enlarging the rock block to be closer to the actual engineering situation. Therefore, the mechanical parameters of the rock mass are determined by combining empirical formula calculation and iterative trial and error of the UDEC-Trigon rock mass model.

[0054] (3) In determining the damage conditions of the roadway anchored rock mass, this invention differs from the existing single inversion uniaxial compression experiment. It considers the actual simulation of roadway roof damage under the influence of mining, assigns confining pressure in the UDEC-Trigon damage evolution model, and inverts the triaxial compression experiment of the rock mass, making the iterative correction of rock mass parameters and the damage process of roof surrounding rock more accurate.

[0055] (4) The method for comprehensively measuring and representing the damage degree of the surrounding rock mass of the coal mining by the fissure development density and the penetration degree in the coal rock mass, comprehensively considers the influence of the confining pressure and the fissure development on the damage variable, and determines the critical instability damage variable Di by the confining pressure 临 As the judgment value, the damage condition in the UDEC-Trigon rock fissure model is monitored in real time.

[0056] (5) The breaking failure criterion of the supporting structure, i.e., the anchor rod or the anchor cable, is added, which is different from the existing invention, and the range of the plastic region of the surrounding rock subjected to shear damage is used as the instability basis in the numerical simulation, the elongation breaking condition of the anchor rod or the anchor cable is visualized, and the roof separation value when the elongation rate of the supporting structure exceeds the elongation rate at which the supporting structure is broken is used as the limit value.

[0057] The technical solutions of the present application are described in further detail below by means of the drawings and the embodiments. BRIEF DESCRIPTION OF DRAWINGS

[0058] Figure 1 The rock stress-strain curve diagram in the embodiment of the present application is shown in the figure.

[0059] Figure 2 The UDEC Trigon model damage evolution process diagram in the embodiment of the present application is shown in the figure.

[0060] Figure 3 The critical damage instability equation Di = f(σ) in the embodiment of the present application is shown in the figure. 临

[0061] Figure 4 The breaking strength test diagram of the anchor rod or the anchor cable in the embodiment of the present application is shown in the figure.

[0062] Figure 5 The longwall working face layout diagram along the strike of the mining face in the embodiment of the present application is shown in the figure.

[0063] Figure 6 The surrounding rock damage fissure evolution diagram in the embodiment of the present application is shown in the figure.

[0064] Figure 7 The stress state diagram of the roadway anchor rod and anchor cable in the embodiment of the present application is shown in the figure.

[0065] Figure 8 The horizontal displacement nephogram of the roadway under different bearing pressure coefficients in the embodiment of the present application is shown in the figure.

[0066] Figure 9 The vertical displacement nephogram of the roadway under different bearing pressure coefficients in the embodiment of the present application is shown in the figure.

[0067] Figure 10 The roadway survey line layout diagram in the embodiment of the present application is shown in the figure.

[0068] ​Figure 11 Figure 1 is a schematic diagram of a roadway roof displacement graph in an embodiment of the present application. DETAILED DESCRIPTION

[0069] The technical solutions of the present application are further described below by means of the accompanying drawings and embodiments.

[0070] Unless otherwise defined, technical terms or scientific terms used in the present application shall have the ordinary meaning understood by a person of ordinary skill in the art to which the present application pertains.

[0071] The terms such as "comprise" or "include" and the like used in the present application mean that the elements before the word are encompassed in the elements listed after the word, and do not exclude the possibility of also encompassing other elements. The terms "in", "out", "up", "down", etc. indicate the orientation or positional relationship shown in the drawings, and are only for the convenience of describing the present application and simplifying the description, and do not indicate or imply that the device or element referred to must have a particular orientation, be constructed and operated in a particular orientation, and therefore cannot be understood as a limitation on the present application. When the absolute position of the described object changes, the relative positional relationship may also change accordingly. In the present application, unless otherwise explicitly specified and limited, the term "attached" and the like should be understood broadly, for example, it can be fixedly connected, or it can be detachably connected, or it can be integrated; it can be directly connected, or it can be indirectly connected through an intermediate medium; it can be the internal connection of two elements or the interaction relationship between two elements. For those of ordinary skill in the art, the specific meaning of the above terms in the present application can be understood according to the specific circumstances.

[0072] EMBODIMENT

[0073] The method for determining the limit value of the roadway roof separation of a weakly cemented mining stratum comprises the following steps:

[0074] Step S1: Obtain the mechanical parameters of the weakly cemented stratum in the mining area of the simulated roadway surrounding rock block through laboratory tests. The mechanical parameters include density, elastic modulus, Poisson's ratio, internal friction angle, cohesion, compressive strength, and tensile strength. The laboratory mechanical parameter determination experiment includes uniaxial compression, Brazilian splitting, and variable angle shear test, and the determination of compressive strength, tensile strength, and shear strength is completed.

[0075] In step S1, the uniaxial compression test uses a cylindrical test piece with a width of 50 mm and a height of 100 mm. The Brazilian splitting, direct shear test, and cylindrical test piece with a diameter of 50 mm are used.

[0076] Step S2: A small model of the same size as the indoor experiment in Step S1 is established using the UDEC Trigon model. Based on the stress-strain curves of the uniaxial compression test and the Brazilian splitting test in Step S1, the contact parameters and block parameters of the small model established using the UDEC Trigon model are inverted to obtain reasonable contact surface parameter values ​​and reasonable block parameter values.

[0077] In step S2, the shear modulus G and bulk modulus K are calculated based on the relationship between the elastic modulus E and Poisson's ratio ν:

[0078]

[0079] In the numerical model, the normal phase stiffness and shear stiffness of the contact surface are derived from the following equations:

[0080]

[0081] In the formula, ΔZ min k is the minimum side length of the block. n For the normal phase stiffness of the contact surface, k s Shear stiffness of the contact surface;

[0082] The method for numerically inverting mechanical parameters using the UDECTrigon model judges the inversion effect by comparing the trends of the indoor test results of the rock sample in step S1 with those of the model-simulated axial stress-strain curves. Figure 1 As shown;

[0083] The contact surface adopts the Residual yield criterion, and the block adopts an elastic model. Scale numerical simulation experiments were conducted on uniaxial compression and Brazilian splitting standard specimens. The parameters were continuously adjusted using a trial-and-error method. The numerical calculation results were compared with the indoor test results through the stress-strain curves of the rock simulation test. When the curve trends are basically consistent, it indicates that the micromechanical parameters of the rock block inverted by the UDECTrigon model are accurate.

[0084] Step S3: Substitute the rock mechanical parameters obtained from the indoor test in Step S1 into the empirical formulas of elasticity theory and strength theory to determine the material parameters in the numerical calculation model, i.e., the rock mass parameters.

[0085] In step S3, the elastic modulus of the rock mass is obtained using the following formula:

[0086]

[0087] Among them, E m E represents the elastic modulus of the rock mass. r Here, represents the elastic modulus of the rock mass, and RQD is the rock quality index. The uniaxial strength of the rock mass is obtained using the following formula:

[0088]

[0089] wherein σ1 is the maximum principal stress at the time of rock mass failure, σ3 is the minimum principal stress at the time of rock mass failure, σ is the uniaxial compressive strength of rock mass, and s is the Hoek constant. ci uniaxial strength of rock mass, m b uniaxial strength of rock mass, m

[0090] Step S4: A UDEC Trigon model is used to establish a uniaxial compression large model with a width-height of 1 m*2 m and a Brazilian splitting large model with a diameter of 1 m, the reasonable contact surface parameter value and the reasonable block parameter Trigon model parameter determined in step S2 are substituted into the large model, and the rock mass mechanical parameter determined in step S2 is used to inverse the contact parameter and the block parameter of the uniaxial compression large model and the Brazilian splitting large model to obtain the rock mass contact surface parameter value and the block parameter value, which specifically include the normal stiffness, the shear stiffness, the internal friction angle, the cohesion, the tensile strength, the bulk modulus, and the shear modulus;

[0091] Step S5: The rationality of the rock mass contact surface parameter value and the block parameter value is verified, a UDEC Trigon module is used to establish a roadway surrounding rock model, the rock mass contact surface parameter value and the block parameter value determined in step S4 are brought into the roadway surrounding rock model for continuous iteration adjustment until the damage degree of the simulated surrounding rock roadway rock mass is consistent with the field situation, and the reasonable block parameter and the contact surface parameter of the model rock mass are determined;

[0092] Step S6: A weakly cemented rock mass damage degree characterization method is established, and specifically, the damage degree of the mining rock mass is comprehensively measured and characterized by the crack development density and the penetration degree in the coal rock mass. According to the occurrence condition of the simulated roadway surrounding rock and the physical and mechanical parameters of the rock stratum in the set range of the simulated roadway, the "damage" Di in the Trigon model of the roadway is defined as the ratio of the contact length (shear crack or tensile crack) of all cumulative failure blocks to the length of all blocks. The rock mass parameters input in step S4 and step S5 are input into the UDEC Trigon model, and the triaxial compression damage-strain relationship curve of the rock mass sample is inversed;

[0093] In step S6, the damage variable Di of the Trigon model is defined as the ratio of the contact length of the failure block to the length of all blocks:

[0094]

[0095] wherein L s is the cumulative tensile crack length at different stages, L n is the cumulative shear crack length at different stages, and L is the cumulative crack length. The triaxial compression damage-strain relationship curve of the rock mass sample inversed in step S6 and the schematic diagram of the damage evolution process of the UDEC Trigon model in step S6 are shown inFigure 2 as shown.

[0096] Step S7: Study the instability mechanism of the roof rock mass; in step S6, the evolution process of the triaxial compression damage-strain relationship curve is divided into four stages, specifically including an initial damage stage in which the damage variable value is small and can be ignored, a damage stable development stage when the stress value reaches 20% to 40% of the peak stress, a damage rapid development stage when the stress value increases to 70% to 90% of the peak stress, and a damage destruction stage in which the damage variable value tends to be stable; when the stress value is 80% to 90% of the peak stress, the damage variable increases slowly and macroscopic cracks appear, which is the instability stage of the roadway roof rock mass.

[0097] Step S8: Simulate the critical value of the roadway roof destruction under mining; in step S6, gradually increase the confining pressure in the UDECTrigon damage evolution model to the confining pressure when the actual roadway roof is mined, and calculate the critical damage variable of the triaxial compression test when the simulation specimen is destroyed; due to the consideration of the influence of the confining pressure, the roadway surrounding rock instability stage in step S6 will occur near 70% to 80% of the peak stress value.

[0098] Step S9: Propose a discrimination method for the instability of the roadway roof surrounding rock; in step S8, fit the changes of the damage variable under different confining pressures, establish a critical damage instability equation Di 临 = f(σ), as shown in Figure 3 , and in the UDECTrigon damage evolution model, monitor the rock cracks under different confining pressures in real time through the critical instability equation, and calculate the damage variable Di value in real time; when the real-time calculated damage variable Di value exceeds the critical damage variable of the roadway instability stage in step S8, the instability is determined.

[0099] Step S10: To ensure the accuracy of the method for calculating the critical value of the roadway roof separation, propose a support structure breakage discrimination method that uses the stress characteristics of the anchoring bearing body support structure to characterize the stability of the roadway roof anchoring bearing body; when the support structure breaks, it means that the subsidence of the anchored rock layer reaches the maximum, that is, the roof separation at this time reaches the limit value; the support structure includes an anchor rod or anchor cable structure;

[0100] Step S11: Optimize the anchor rod or anchor cable structure unit; the mechanical parameters of the optimized support unit of the UDECTrigon model block body are shown in Tables 1, 2 and 3, and the breaking strength test of the anchor rod or anchor cable is shown in Figure 4 , and the calculated tensile strength of the anchor cable is 390 kN and the shear strength of the anchor rod is 100 kN.

[0101] Table 1

[0102]

[0103] Table 2

[0104] anchoring section Shear modulus / GPa Yield force / kN free section 0 0 anchoring section 20 400

[0105] Table 3

[0106] anchoring section Shear modulus / GPa Yield force / kN free section 0 0 anchoring section 20 200

[0107] In step S11, the optimization of the anchor rod or anchor cable structure unit specifically includes the following steps:

[0108] S111: On the basis of the existing Cable unit of UDEC, the anchor cable breakage failure criterion U(i) ≥ Umax(i) is added, U(i) representing the length of any unit of the anchor cable, which is detected by the existing module of UDEC, and Umax(i) being the limit length of the corresponding unit, which is artificially set;

[0109] S112: The Rockbolt structure unit is optimized, and the Rockbolt failure breakage criterion ε pl ≥ tfstrain; in the formula, tfstrain is artificially set; ε pl is the total tensile yield strain of any unit of the anchor rod, which is detected by the existing module of UDEC:

[0110]

[0111] In the formula, is the axial strain, d is the diameter of the anchor rod, θ pl is the average rotation angle of the component, and L is the length of the component;

[0112] If the total tensile yield strain ε pl of any unit of the Rockbolt element of the anchor rod exceeds the artificially set tensile failure strain limit tfstrain, the stress and bending moment of the anchor rod component will instantaneously decrease to 0, and it is considered that the anchor rod has broken at this place. The anchor cable is simulated by the Cable unit, and the anchor rod is simulated by the Rockbolt unit.

[0113] Step S12: A UDEC Trigon model of the working face recovery roadway in the engineering practice is established, the boundary conditions of the calculation model are input, and the corrected rock mass mechanical parameters are assigned as the rock layer mechanical parameters in the model. The block is simulated by an elastic model, and the contact surface is simulated by a Mohr-Coulomb model. The determination of the mechanical parameters of the block of the UDEC Trigon model is based on the plastic mechanics theory, and the corresponding constitutive relationship is established. At the same time, the geological conditions of the coal mine also need to be considered, and the influence of the model boundary on the numerical simulation is considered. A 15m boundary is left on the roof and floor of the roadway, and a 15m boundary is left on the left and right of the roadway. The block parameters and the contact surface parameters of the UDEC Trigon model are shown in Table 4.

[0114] Table 4

[0115]

[0116]

[0117] Step S13: Establishing the roof separation simulation scheme and the separation limit value criterion: establishing the model and assigning the parameters → applying the stress boundary conditions according to different mining conditions → calculating the original rock stress → mining the roadway to the stable; in step S13, the roof separation under different stress concentration conditions is simulated, as shown in Table 5, and the roof separation values inside and outside the anchoring segment are outputted, when the elongation rate of the anchor rod exceeds the breaking elongation rate, it indicates that the anchoring material is damaged at this time, and the roof separation value at this time is the limit value.

[0118] The roadway instability condition of the UDECTrigon model block is that the roadway surrounding rock loose circle is greater than the anchoring range, and the elongation rate of the anchor rod exceeds the breaking elongation rate, that is, the critical value of the roof separation is calculated according to the breaking of the anchoring material (anchor rod and anchor cable), and further, the sinking value of the limit roadway roof when the anchor rod and anchor cable break is the critical value of the roof separation.

[0119] Table 5

[0120]

[0121] Step S15: According to the UDECTrigon model, the surrounding rock response law of the anchor rod and anchor cable support mining roadway under the influence of mining is simulated, different support pressure coefficients K are assigned, and the greater the K value represents the greater the applied pressure, the surrounding fissure field, displacement field and stress characteristics of the anchor rod or anchor cable of the anchor net and cable supported roadway are simulated;

[0122] The surrounding rock roadway established by the UDECTrigon model will be affected by the mining stress in the mining process, the support pressure in the roadway surrounding rock gradually increases, and the support pressure in front of the working face is the largest, therefore, the roof simulation scheme should take different roof strength, roadway width and mining stress concentration coefficient, as shown in Table 6, further, the surrounding rock response law under the influence of mining is simulated by taking the siltstone, respectively taking the support pressure coefficient K = 1.0, 1.5, 2.0, 2.5, and the surrounding fissure field, displacement field and stress characteristics of the anchor rod or anchor cable of the anchor net and cable supported roadway are obtained.

[0123] Table 6

[0124]

[0125] The damage fissure of the roadway, the stress state of the anchor rod and anchor cable, the horizontal displacement nephogram and the vertical displacement nephogram under different support pressure coefficients are respectively shown in Figures 6-9 and Figure 10The displacement data in the measuring lines arranged on the roadway roof are extracted to obtain the surface displacement and deep displacement of the roadway roof.

[0126] The roadway damage fissure, stress state of the anchor rod and anchor cable are shown in Figure 6 and Figure 7 With the increase of the support pressure coefficient, the damage fissure around the roadway expands continuously, and the roof part of the roadway surrounding rock appears to be damaged. For example, when K = 1, the tensile fissure is generated in the shallow part of the roadway roof, the stress of the roadway anchor rod and anchor cable is uniform, and no breakage phenomenon occurs, which indicates that the current support mode can well protect the roadway when it is not affected by mining. When K = 2.5 or more than 2.5, the tensile fissure in the roadway roof gradually develops to the deep part, the shallow tensile fissure develops significantly and penetrates each other, the surrounding rock loose zone exceeds most of the anchoring depth of the anchor rod, and most of the anchor rod is broken, and the stability of the roadway gradually deteriorates.

[0127] The displacement nephogram is shown in Figure 8 and Figure 9 With the continuous increase of the support pressure coefficient, the displacement around the roadway continuously increases, and the maximum deformation area is located in the roof. The overall deformation shows that the roof deformation is greater than the deformation of the two sides. With the increase of K value, the circular contour lines of the roadway sides and the roof and floor gradually fuse, and the roof displacement increment is greater than the displacement increment of the two sides. After the K value increases from 1.5 to 2.5, the roof subsidence increases from 123 mm to 349 mm.

[0128] Step S16: Measuring lines are arranged on both sides of the roadway model, the roof and floor in the UDEC simulation, and the displacement data in the measuring lines are extracted to obtain the surface displacement and deep displacement of the roof and floor of the roadway;

[0129] The arranged red observation lines are shown in Figure 10 The displacement values of the roadway surrounding rock unit on the observation lines are extracted, and the extracted data are plotted as the evolution curves of the crossheading deep displacement and surface displacement shown in Figure 11 With the continuous increase of the support pressure coefficient, the roof displacement slowly increases at the depth of 10 m to 4 m, and sharply increases within the depth of 4 m. With the increase of the support pressure coefficient, the change rate of the displacement increases. When K = 2.5, the surface displacement of the roadway roof reaches the maximum value, i.e. 349 mm. It can be known from Figures 4-10 that the roof deformation presents overall subsidence. With the continuous increase of the support pressure coefficient, the relative increase of the displacement of the anchoring segment of the surrounding rock is obvious (113 mm). The separation of the anchoring segment of the roof anchor rod is 61 mm, and the separation of the anchoring segment of the anchor cable is 52 mm. The anchoring segments of the roof anchor rod and anchor cable are not broken, and most of the anchor rod in the side part is broken, which indicates that the support body of the roof anchor rod or anchor cable does not have structural damage.

[0130] Step S17: the surrounding rock response law of the anchor rod and anchor supporting mining roadway under the influence of mining is simulated based on the UDEC Trigon model, the surrounding rock deformation characteristics of the working face mining roadway are obtained, the anchor rod anchoring segment separation limit value and the anchor cable anchoring segment separation limit value when the roadway is unstable are obtained from the roadway instability condition, the separation limit value of the roadway roof is the sum of the anchor rod anchoring segment separation limit value and the anchor cable anchoring segment separation limit value, and the corresponding anchor rod anchoring segment separation and anchor cable anchoring segment separation critical values are obtained by taking a reasonable safety factor.

[0131] Based on the dynamic response law of the thick immediate roof supporting surrounding rock roadway of the UDEC model block, the surrounding rock deformation characteristics of the first mining working face mining roadway can be obtained. According to Tables 1-3, when the anchor rod anchoring segment separation limit value is 61mm and the anchor cable anchoring segment separation limit value is 52mm, the surrounding rock is unstable, the anchor rod anchoring segment separation limit value is taken as 1 / 2 of the roof anchor rod anchoring segment critical value, i.e. 30.5mm, the anchor cable anchoring segment separation limit value is taken as 1 / 2 of the anchor cable anchoring segment critical value, i.e. 26mm, and the separation indicator cannot monitor whether it is the anchor cable segment separation, so the roof cumulative separation critical value is 56.5mm (30.5+26=56.5mm), i.e. the roof subsidence limit is 56.5mm.

[0132] Therefore, compared with the theoretical calculation result, the accuracy of the roadway roof separation limit value calculation and monitoring is improved by 34.78% by using the above-mentioned weakly cemented mining stratum roadway roof separation limit value determination method.

[0133] Finally, it should be noted that: the above examples are only used to illustrate the technical solutions of the present application and not to limit them, although the present application has been described in detail with reference to the preferred embodiments, those skilled in the art should understand that the technical solutions of the present application can still be modified or replaced by equivalents, and these modifications or equivalent replacements cannot make the modified technical solutions deviate from the spirit and scope of the technical solutions of the present application.

Claims

1. A method for determining the roof separation limit value of a roadway in a weakly cemented mining stratum, characterized in that: Comprise the following steps: Step S1: Obtain the mechanical parameters of the weakly cemented surrounding rock mass in the roadway in the simulated weakly cemented stratum through indoor tests, the mechanical parameters including density, elastic modulus, Poisson's ratio, internal friction angle, cohesion, compressive strength and tensile strength, the indoor mechanical parameter determination experiment including uniaxial compression, Brazilian splitting and variable angle shear test, and the determination of compressive strength, tensile strength and shear strength is completed; Step S2: A small model of the same size as the indoor experiment in step S1 is established by using the UDEC Trigon model, and the contact parameters and block parameters of the small model established by using the UDEC Trigon model are inversed to obtain reasonable contact surface parameter values and reasonable block parameter values according to the stress-strain curves of the uniaxial compression test and the Brazilian splitting test in step S1; Step S3: The rock mass mechanical parameters obtained through the indoor test of step S1 are substituted into the empirical formula of the elastic theory and the strength theory to determine the mechanical parameters of the rock mass; Step S4: A uniaxial compression large model with a width of 1m and a height of 2m and a Brazilian splitting large model with a diameter of 1m are established by using the UDEC Trigon model, the reasonable contact surface parameter values and reasonable block parameter Trigon model parameters determined in step S2 are substituted into the large model, and the rock mass mechanical parameters determined in step S2 are inversed to the contact parameters and block parameters of the uniaxial compression large model and the Brazilian splitting large model to obtain the rock mass contact surface parameter values and block parameter values; Step S5: Verify the rationality of the rock mass contact surface parameter values and block parameter values, establish a numerical model of the stability of the surrounding rock of the roadway by using the UDEC Trigon module, and continuously iterate and adjust the numerical model of the stability of the surrounding rock of the roadway by bringing the rock mass contact surface parameter values and block parameter values determined in step S4 into the numerical model until the damage degree of the simulated surrounding rock of the roadway is consistent with the engineering site to determine the reasonable block parameter and contact surface parameter of the model rock mass; Step S6: Establish a weakly cemented rock mass damage degree characterization method, input the rock mass block parameter and contact surface parameter of the roadway surrounding rock model after iteration and trial correction in step S5 into the UDEC Trigon model, and inverse the damage-strain relationship curve of the rock mass sample under triaxial compression state; Step S7: Study the instability mechanism of the weakly cemented surrounding rock roadway roof rock mass; Step S8: Simulate the critical damage value of the weakly cemented surrounding rock roadway roof anchoring load carrier under the influence of mining; Step S9: Propose a damage discrimination method for the instability of the weakly cemented surrounding rock roadway roof anchoring load carrier; Step S10: To ensure the accuracy of the method for calculating the critical value of the roadway roof separation, a support structure breakage discrimination method is proposed to characterize the stability of the roadway roof anchoring load carrier by the stress characteristics of the support structure, and when the support structure breaks, it means that the subsidence of the anchored rock layer reaches the limit, and the support structure includes anchor rod or anchor cable; Step S11: Optimize the anchor rod or anchor cable structure unit; Step S12: Establish a UDEC Trigon model of the actual engineering condition mining roadway, input the boundary conditions of the calculation model, and take the assigned and corrected rock mass mechanical parameters as the rock mass mechanical parameters in the model, the block adopts an elastic model, and the contact surface adopts a Mohr-Coulomb model; Step S13: Establishing the roof separation simulation scheme and the separation limit value criterion: establishing the model and assigning parameters → applying stress boundary conditions according to different mining conditions → calculating the original rock stress → excavating the roadway to stability; Step S14: According to the UDEC Trigon model, the surrounding rock response law of the anchor rod and anchor cable support mining roadway under the influence of mining is simulated, and different support pressure coefficients K are assigned, the larger the K value, the greater the pressure applied, and the surrounding fissure field, displacement field and stress characteristics of the anchor rod or anchor cable of the anchor net support roadway are simulated; Step S15: Displacement data is extracted from the displacement data of the measuring line arranged on both sides of the roadway model and the roof and floor to obtain the surface displacement and deep displacement of the roadway roof and floor; Step S16: Based on the UDEC Trigon model, the surrounding rock response law of the anchor rod and anchor cable support mining roadway under the influence of mining is simulated to obtain the deformation characteristics of the working face mining roadway, and the separation limit values of the anchor rod anchoring section and the anchor cable anchoring section at the time of instability are obtained from the instability conditions of the roadway, the separation limit value of the roof of the roadway is the sum of the separation limit values of the anchor rod anchoring section and the anchor cable anchoring section, and the corresponding critical values of the anchor rod anchoring section separation and the anchor cable anchoring section separation are obtained by taking a reasonable safety factor.

2. The method for determining the roof separation limit value of the weakly cemented mining stratum roadway according to claim 1, characterized in that: In step S1, the uniaxial compression test uses a cylindrical specimen with a width of 50 mm and a height of 100 mm, and the Brazilian split and variable-angle shear tests use a cylindrical specimen with a diameter of 50 mm.

3. The method for determining the roof separation limit value of the weakly cemented mining stratum roadway according to claim 2, characterized in that: In step S2, the shear modulus G and the bulk modulus K of the numerical model are calculated according to the relationship between the elastic modulus E and the Poisson's ratio v in S1: The normal stiffness and shear stiffness of the contact surface in the numerical model are obtained from the following equations: In the formula, ΔZ min k is the minimum side length of the block. n For the normal phase stiffness of the contact surface, k s Shear stiffness of the contact surface; The UDEC Trigon model numerical inversion mechanics parameter method judges the inversion effect by whether the variation trend of the axial stress-strain curve of the rock block sample in the laboratory test and the model simulation in step S1 is similar; The contact surface uses the Residual yield criterion, and the block uses an elastic model to perform a numerical simulation experiment on the uniaxial compression and Brazilian split standard specimens, and the parameters are continuously adjusted using the trial-and-error method. The numerical calculation results are compared with the laboratory test results through the stress-strain curve of the rock simulation test, and when the curve trend is basically consistent, the rock micro-mechanics parameters inverted by the UDEC Trigon model are accurate.

4. The method for determining the roof separation limit value of the weakly cemented mining stratum roadway according to claim 3, characterized in that: In step S3, the rock elastic modulus is obtained by the following formula: where E m is the rock mass elastic modulus, E r is the rock block elastic modulus, and RQD is the rock quality index. The rock mass uniaxial strength is obtained by the following formula: where σ1is the maximum principal stress at failure of the rock mass, σ3is the minimum principal stress at failure of the rock mass, σ ci Uniaxial strength of rock mass, m b s is the Hoek constant.

5. The method for determining the roof separation limit value of the weakly cemented mining stratum roadway according to claim 4, characterized in that: In step S6, the Trigon model damage variable Di is defined as the ratio of the failed block contact length to the length of all blocks: where L s is the cumulative tensile fracture length at different stages, L n is the cumulative shear fracture length at different stages; L is the cumulative fracture length.

6. The method for determining the roof separation limit value of the weakly cemented mining stratum roadway according to claim 5, characterized in that: In step S7, the evolution process is divided into four stages, including the initial damage stage where the damage variable value is small and can be ignored, the damage stable development stage when the stress value reaches 20% to 40% of the peak stress, the damage rapid development stage when the stress value increases to 70% to 90% of the peak stress, and the damage failure stage when the damage variable value tends to be stable. When the stress value is 80% to 90% of the peak stress, the damage variable increases slowly and macro cracks appear, which is the instability stage of the rock mass.

7. The method for determining the weakly cemented mining stratum roadway roof abscission layer limit value according to claim 6, characterized in that: In step S8, the confining pressure is gradually increased to the confining pressure of the real roadway roof mining in the UDEC Trigon damage evolution model, and the critical damage variable of the triaxial compression test when the simulation specimen is damaged is calculated respectively, and the stress value is near 70% to 80% of the peak stress.

8. The method for determining the weakly cemented mining stratum roadway roof abscission layer limit value according to claim 7, characterized in that: In step S8, the change of damage variable under different confining pressures is fitted to establish the critical damage instability equation Di 临 = f(σ), and the rock mass crack under different confining pressures is monitored in real time by the critical instability equation in the UDEC Trigon damage evolution model, and the damage variable Di value at this time is calculated in real time. When the real-time calculated damage variable Di value exceeds the critical damage variable in the rock mass instability stage in step S7, instability is determined.

9. The method for determining the roof separation limit value of the weakly cemented mining stratum roadway according to claim 8, characterized in that: In step S11, the optimization of the anchor rod or anchor cable structure unit specifically includes the following steps: S111: On the basis of the existing Cable unit of UDEC, the anchor cable breaking failure criterion U(i) ≥ Umax(i) is added, U(i) represents the length of any unit of the anchor cable, which is obtained by detecting the existing module of UDEC, and Umax(i) is the limit length of the corresponding unit, which is artificially set; S112: Optimizing Rockbolt structural unit, Rockbolt failure breaking criterion ε pl ≥ tfstrain; in the formula, tfstrain is artificially set; ε pl It is the total tensile yield strain of the anchor rod unit, which is detected by the existing module of UDEC: wherein is the axial strain, d is the anchor diameter, θ pl is the average rotation of the member, L is the length of the member; Rockbolt element total tensile yield strain ε pl When the tensile failure strain limit tfstrain is exceeded, the stress and bending moment of the rock bolt member instantaneously drop to zero, and the rock bolt is considered to have failed at this point. The cable element is used to simulate the cable and the rock bolt element is used to simulate the rock bolt in the model.

10. The method for determining the roof separation limit value of the weakly cemented mining stratum roadway according to claim 9, characterized in that: In step S16, the roof separation values inside and outside the anchoring section are output by simulating the roof separation under different stress concentration conditions. When the elongation of the anchor rod exceeds the elongation at break, it indicates that the anchoring material is damaged or the damage variable Di of the surrounding rock in the anchoring range of the support structure exceeds the critical damage value Di 临 At this time, the roof separation value is the limit value.

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