Non-through pre-splitting roof cutting method for overlying strata of fully mechanized caving face of coal mine

Abstract

The invention provides a coal mine underground driving working face outburst prediction method, which comprises the following steps of S1, arranging drill holes at equal intervals along the inclination direction of a working face, and forming a top cutting angle between the drill holes and the working face; s2, establishing a numerical model of the non-through pre-splitting top cutting surface of the roadway roof in FLAC3D software; s3, N groups of drilling hole top cutting angles are set in the numerical model for non-through presplitting blasting effect simulation, and the needed drilling hole and working face top cutting angle is selected; s4, a pre-generated crack is manufactured in the overlying strata key layer at the drilling position in a blasting mode, and an advanced non-through pre-splitting top cutting face is constructed; s5, the length of a pre-generated fracture formed in the overlying strata key layer is calculated, and the drill hole interval is determined according to the length of the pre-generated fracture; and S6, obtaining a numerical model, obtaining an elastic strain energy distribution cloud picture of the overlying rock key layer, including an elastic strain energy distribution cloud picture of the key layer when the top is not cut and an elastic strain energy distribution cloud picture of the key layer of N groups of different drilling top cutting angles, and obtaining a required pre-splitting top cutting distance.

Description

Technical Field

[0001] The invention relates to the technical field of coal mine rock burst prevention and control, and in particular to a method for non-penetrating pre-splitting and top cutting of overburden in a fully-mechanized caving working face in a coal mine. Background Art

[0002] With the continuous exploitation of coal resources, the safety issues of deep mines have become increasingly prominent. Deep mine mining not only faces challenges such as deep coal seams and severe mining disturbances, but also the intensity and frequency of disasters such as rock bursts have increased, posing a serious threat to the safety of miners and the sustainable development of mines. How to effectively prevent and control rock bursts in fully mechanized caving of thick coal seams has become an important issue that must be solved for safe and efficient mining of deep thick coal seams.

[0003] In the fully mechanized top coal mining of thick coal seams, the overlying rock strata are very likely to form dynamic loads due to fracture instability, leading to rock burst accidents on the working face. At present, in order to solve the impact danger caused by large-scale movement of the roof during fully mechanized top coal mining, domestic and foreign scholars often use deep hole blasting to pre-split the rock strata to reduce the intensity of roof movement. However, there are also some urgent problems to be solved in the application of deep hole blasting technology, such as insufficient accurate grasp of the blasting layer and too dense blasting charges. These problems have led to an increase in economic investment and a decrease in charging efficiency, affecting the economic benefits and application effects of deep hole blasting technology. Summary of the invention

[0004] In order to solve the problems existing in the above-mentioned prior art, the present invention provides a method for non-penetrating pre-splitting and top cutting of overburden in a fully-mechanized caving working face in a coal mine, including the following specific technical solutions.

[0005] A method for non-penetrating pre-splitting and top cutting of overburden in a fully-mechanized caving working face of a coal mine comprises the following steps:

[0006] S1, arrange the drill holes at equal intervals along the inclination of the working surface, and make the drill holes form a cutting angle with the working surface;

[0007] S2, in FLAC 3D The numerical model of the non-penetrating pre-cracked top surface of the tunnel roof is established in the software, which specifically includes the following steps:

[0008] S201, constructing a numerical model according to the actual engineering dimensions of the mine to be measured;

[0009] S202, set the number of overlying rock layers in the numerical model, set the overlying rock layer to the Mohr-Coulomb elastoplastic material model, and 3D The corresponding unit module of the software assigns the rock physical and mechanical parameters corresponding to the overlying rock layer;

[0010] S203, using FLAC 3DThe software’s meshing tool completes the meshing of the numerical model;

[0011] S204, according to the pre-splitting design scheme in the actual project, defining the starting point, the end point and the depth of the pre-splitting surface in the numerical model, so as to realize the construction of the non-through pre-splitting top cutting structure;

[0012] S205, setting boundary conditions of the numerical model, specifically: setting a fixed boundary at the bottom of the numerical model and setting a displacement constraint in the vertical direction; setting a displacement constraint in the horizontal direction on the side of the numerical model according to the actual engineering situation;

[0013] S206, Run FLAC 3D The software performs initial equilibrium calculations to allow the numerical model to reach equilibrium under the initial stress state;

[0014] S3, setting N groups of drilling top cutting angles in the numerical model to simulate the effect of non-penetrating pre-splitting blasting, and selecting the required drilling top cutting angles of the drilling hole and the working surface;

[0015] S4, according to the drilling top cutting angle obtained in step S3, using blasting method at the drilling location to create pre-existing cracks in the overburden key layer, and constructing an advanced non-penetrating pre-crack cutting top surface;

[0016] S5, calculating the length of pre-existing cracks formed in the key overburden layer, and determining the drilling spacing according to the length of the pre-existing cracks;

[0017] S6. According to the numerical model of the non-penetrating pre-splitting top cutting surface of the tunnel roof established in step S2, a cloud map of the elastic strain energy distribution of the key layer of the overburden is obtained, including a cloud map of the elastic strain energy distribution of the key layer before the top cutting and a cloud map of the elastic strain energy distribution of the key layer of N groups with different drilling top cutting angles; a relationship diagram between the maximum elastic strain energy and the working face advancement distance is prepared according to the cloud map of the elastic strain energy distribution of the key layer of the overburden, so as to obtain the required pre-splitting top cutting spacing.

[0018] Furthermore, the minimum cutting angle formed by the drill hole and the working surface is calculated by the following formula:

[0019]

[0020] In the formula,

[0021] a is the length of the pre-existing crack, m; K c is the fracture toughness; λ is the crack extension compression-shear ratio coefficient; π is the circumference of the circumference, which is taken as 3.14; F M(a / b) is the bending force corresponding to the bending moment of the key layer, N; q is the uniformly distributed load of the cantilever beam; l is the length of the basic top lateral cantilever beam, m; θ is the top cutting angle; b1 is the thickness of the key layer, m.

[0022] Further, in step S203, through FLAC3D The numerical model is meshed by the meshing module of the software, or by adaptive meshing, manual meshing, or increasing the mesh density.

[0023] Further, in step S204, by 3D The contact surface unit is set in the software to simulate the mechanical properties of the pre-crack surface; the strength parameter ratio of the damaged rock mass to the undamaged rock mass is set as: c / σ c0 =0.28;σ t / σ t0 =0.41; E / E0=0.625; c / c0=0.54;where σ c is the uniaxial compressive strength of damaged rock mass, Pa; σ c0 is the uniaxial compressive strength of undamaged rock mass, Pa; σ t is the uniaxial tensile strength of the damaged rock mass, Pa; σ t0 is the uniaxial tensile strength of the undamaged rock mass, Pa; E is the elastic modulus of the damaged rock mass, Pa; E0 is the elastic modulus of the undamaged rock mass, Pa; is the friction angle in the damaged rock mass, °; is the friction angle of the undamaged rock mass, °; c is the cohesion of the damaged rock mass, Pa; c0 is the cohesion of the undamaged rock mass, Pa.

[0024] Furthermore, in step S3, three groups of drilling top cutting angles are set to simulate the non-penetrating pre-splitting blasting effect, specifically: the first group of 50°, 65°, 80° combination angles, the second group of 45°, 60°, 75° combination angles, and the third group of 40°, 55°, 70° combination angles; and key layer elastic strain energy distribution cloud maps are made for the three groups of drilling top cutting angles.

[0025] Further, 45°, 60° and 75° combined fan-shaped holes are set as the required cutting angles between the drilling hole and the working surface.

[0026] Further, in step S4, in order to form non-through pre-existing cracks in the key layer of the overburden without causing through cracks, the radius of the crushing zone and the radius of the crushing zone generated after the explosion of the explosive are calculated by the following steps: S401, the strong impact load of the columnar charge under the condition of uncoupled charge is calculated by the following formula:

[0027]

[0028] In the formula,

[0029] P is the initial pressure of the shock wave in the rock mass, MPa; ρ0 is the density of the explosive, kg / m 3 ;D vis the detonation velocity of the explosive, m / s; K is the charge uncoupling coefficient; η is the adiabatic index of the explosion product expansion; l c is the axial coefficient of the charge; n is the pressure increase coefficient when the explosion product expands and collides with the blast hole wall;

[0030] S402, calculate the dynamic Poisson's ratio of the rock mass by the following formula:

[0031] μ d =0.8μ

[0032] In the formula, μ d is the dynamic Poisson's ratio of the rock mass, and μ is the static Poisson's ratio of the rock mass;

[0033] S403, calculate the stress intensity at any point in the rock mass using the following formula:

[0034]

[0035] In the formula, σ i is the stress intensity at any point in the rock mass; σ r is the radial stress in the rock mass, MPa; b2 is the lateral stress coefficient, μ d is the dynamic Poisson's ratio of the rock mass; σ r is the radial stress of any particle in the rock mass; p is the stress of the rock mass before being locally disturbed, Pa; is the ratio distance, r is the distance from the calculation point to the charge center, m; r b is the radius of the blasthole, m; a is the attenuation index of load propagation, The positive and negative signs represent the shock wave area and stress wave area respectively; μ d is the dynamic Poisson's ratio of the rock mass;

[0036] S404, the conditions for rock mass failure are identified by the following formula:

[0037] σ i ≥σ0

[0038]

[0039] Where:

[0040] σ i is the stress intensity at any point in the rock mass; σ0 is the uniaxial stress failure strength of the rock, MPa; σ cd is the uniaxial compressive strength of rock, MPa; σ td is the uniaxial tensile strength of rock, MPa;

[0041] S405, based on steps S401-S404, the radius of the crushing zone and the radius of the crushing zone are calculated respectively by the following formulas:

[0042] Crushing zone radius:

[0043]

[0044] Crushing zone radius:

[0045]

[0046] Where:

[0047]

[0048] R c is the radius of the rock mass crushing zone, m; R p is the radius of the rock mass crushing zone, m; ρ0 is the density of the explosive, kg / m 3 ;D v is the detonation velocity of the explosive, m / s; n is the pressure increase coefficient when the explosion product expands and collides with the blast hole wall, which is 10; K is the charge decoupling coefficient, which is 1.5-2.5; l c is the axial coefficient of the charge, which takes the value of 1; B is the comprehensive coefficient related to the lateral stress coefficient of the rock and the dynamic Poisson's ratio, with no specific unit; a is the length of the pre-existing crack, m; r b is the blasthole radius, m; σ cd is the uniaxial compressive strength of rock, MPa; σ R is the radial stress on the interface between the crushing zone and the crushing zone, MPa, b2 is the lateral stress coefficient; β is an index related to material crushing characteristics, stress distribution, etc., and has no specific unit; μ d is the dynamic Poisson's ratio of the rock mass; σ td is the uniaxial tensile strength of rock, MPa.

[0049] Furthermore, in step S5, the length of the pre-existing crack is calculated by the following formula:

[0050]

[0051] In the formula,

[0052] K C =λ∑K Ⅰ +|∑K Ⅱ |

[0053]

[0054] P=ρgH

[0055] a is the length of the pre-existing crack; K Cis the fracture toughness; λ is the crack extension compression-shear ratio coefficient; π is the circumference of a circle, which is taken as 3.14; F M(a / b) is the bending force corresponding to the bending moment of the key layer, N; q is the uniform load of the cantilever beam; l is the length of the basic top lateral cantilever beam, m; θ is the pre-crack cutting top angle; b1 is the thickness of the key layer, m; F τ(a / b) is the shear force of the key layer, N; F σ(a / b) is the geometric correction parameter; T is the horizontal extrusion pressure on both sides of the overlying rock layer, N; K Ⅰ is the stress intensity factor at the crack tip of mode I; K Ⅱ is the stress intensity factor at the crack tip of mode II; l' is the periodic pressure step, m; S0 is the length of the working surface, m; P is the overburden load; ρ is the total average density of the overburden, kg / m 3 ; g is the acceleration due to gravity, 9.8m 2 / s; H is the vertical height of the overlying rock layer, m.

[0056] Furthermore, in step S6, the pre-splitting top cutting distance is set to 10m.

[0057] Furthermore, the feasibility of constructing non-penetrating pre-splitting and cutting of the tunnel roof overburden is verified through the following steps:

[0058] Step 1: Calculate the energy V released during the breaking of the tunnel roof using the following formula:

[0059]

[0060] In the formula,

[0061] V is the energy released during the breaking process of the tunnel roof, J; V G is the gravitational potential energy generated by the sinking of the overburden, J; V ε is the bending strain energy of the cantilever beam at the critical breaking state, N; Q is the load of the overlying rock stratum on the cantilever beam, N; h s is the tunnel roof sinking height, m; M max is the maximum cantilever beam moment, N·m; l1 is the critical breaking length of the suspended roof in the goaf, m; E is the elastic modulus of the cantilever beam, Pa; I is the moment of inertia of the cantilever beam section, m 4 ; q is the uniformly distributed load of the cantilever beam, N / m.

[0062] Step 2: Based on the energy released during the breaking of the tunnel roof, the energy release-roof sinking height relationship diagram, the energy release-overburden load relationship diagram, and the energy release-cantilever beam length relationship diagram are drawn respectively, and the above relationship diagrams are concluded: In order to prevent and control impact ground pressure, cracks can be set in the key layer of the overburden, and a non-penetrating pre-crack cutting top of the tunnel roof overburden is constructed.

[0063] Based on the above technical solution, the present invention has the following beneficial effects:

[0064] The method of the present invention, by recording the non-penetrating pre-splitting and top cutting technology, reveals the mechanical mechanism of non-penetrating pre-splitting and top cutting of key layers in fully mechanized caving mining. By manufacturing the pre-splitting surface in the key layer, the length of the cantilever beam and the stress concentration can be effectively reduced, thereby controlling the release of energy and reducing the risk of rock burst; the spatial parameter values ​​of the non-penetrating pre-splitting slits are mastered, providing a theoretical basis for technical implementation; a three-dimensional numerical model of non-penetrating pre-splitting and top cutting is established, and it is determined that the effect is most significant when the fan-shaped holes are combined with the top cutting angles of 45°, 60° and 75° and the pre-splitting and top cutting spacing is 10m. It can effectively reduce the probability of rock burst and promote the safety and efficiency of pre-splitting blasting and top cutting. BRIEF DESCRIPTION OF THE DRAWINGS

[0065] Figure 1 : Schematic diagram of the relationship between energy release and roof sinking height;

[0066] Figure 2 : Schematic diagram of energy release-overburden load relationship;

[0067] Figure 3 : Schematic diagram of the relationship between energy release and cantilever beam length;

[0068] Figure 4 : Schematic diagram of the mechanical model for calculating the pre-crack cutting top angle;

[0069] Figure 5 : Schematic representation of the mechanical parameters of the geological body of the overlying rock layer on the tunnel roof;

[0070] Figure 6 : Schematic diagram of mechanical parameters of rock mass under blasting damage;

[0071] Figure 7 : Schematic diagram of the numerical model of non-through pre-crack cutting top surface of the tunnel roof;

[0072] Figure 8 : Schematic diagram of the elastic strain energy distribution cloud diagram of the key layer when the top is not cut;

[0073] Fig. 9 : Schematic diagram of the distribution cloud diagram of elastic strain energy of the key layer when the top is cut at a combination angle of 50°, 65°, and 80°;

[0074] Fig.10 : Schematic diagram of the distribution of elastic strain energy of the key layer when the top is cut at a combination angle of 45°, 60°, and 75°

[0075] Fig.11 : Schematic diagram of the distribution cloud diagram of elastic strain energy of the key layer when the top is cut at a combination angle of 40°, 55°, and 70°;

[0076] Fig.12 : Schematic diagram of the relationship between the elastic strain energy of the key layer and the advancement distance of the working face;

[0077] Fig.13 : Schematic diagram of the relationship between the maximum elastic strain energy and the working face advancement distance;

[0078] Fig.14 : Schematic diagram of energy release efficiency of key roof layers and working face advancement distance under different top cutting spacing;

[0079] Fig.15 : Schematic diagram of the 6310 working surface layout;

[0080] Fig.16 : Schematic diagram of the layout of blasting and top drilling holes in the 6310 working face;

[0081] Fig.17 : Schematic diagram of rock blasting damage zones;

[0082] Fig.18 :Schematic diagram of the total frequency curve of microseismic events in a single day;

[0083] Fig.19 :Schematic diagram of the total energy curve of microseismic events in a single day. DETAILED DESCRIPTION

[0084] It should be noted that:

[0085] 1. Certain words are used in the specification and claims to refer to specific components. Those skilled in the art should understand that technicians may use different nouns to refer to the same component. This specification and claims do not use the difference in nouns as a way to distinguish components, but use the difference in the functions of the components as the criterion for distinction. Unless otherwise defined, the technical terms or scientific terms used in this disclosure should have the common meaning understood by people with ordinary skills in the field to which this disclosure belongs.

[0086] 2. The key layer of overburden in coal mine tunnel refers to the rock layer that plays a major role in controlling the movement of overburden and the appearance of mine pressure during coal mining. The deflection of these rock layers is smaller than that of the underlying rock layers during deformation, and they do not deform in coordination with the underlying rock layers. Before breaking, the key layer bears the weight of the overburden in the form of a plate structure or a beam structure. The identification or calculation of the key layer is a commonly used technology in the field of coal mining, and will not be repeated in this embodiment.

[0087] 3. The fracture of the key layer of the overburden rock in the tunnel roof is elastic-brittle fracture, and the fracture is subject to the maximum circumferential stress criterion. The crack size generated in the key layer of the overburden rock by blasting or hydraulic fracturing is much smaller than the area of ​​the roof of the tunnel working face. After the tunnel working face is mined, the cracks in the middle of the roof are easy to collapse, but a certain area of ​​hanging plate is left in the upper and lower end areas without collapse. Therefore, in this embodiment, the pre-cracks in the key layer of the overburden rock are simplified to the cantilever beam crack expansion of the key layer under the pre-crack cutting condition for processing.

[0088] 4. The suspended rock layer refers to the roof rock layer above the goaf that is temporarily suspended in the process of coal mining without collapsing in time. The suspended rock layer is part of the overlying rock layer.

[0089] This embodiment describes a method for non-penetrating pre-splitting and top cutting of overburden in a fully mechanized caving working face in a coal mine, comprising the following steps:

[0090] S1, feasibility verification of pre-setting cracks in the key layer of the tunnel roof overburden and constructing non-penetrating pre-crack cutting of the tunnel roof overburden, including the following steps:

[0091] S101, the energy released during the roadway roof rupture is the main energy source leading to rock burst events. The released energy mainly includes: bending strain energy V released by the roadway roof rupture ε , the gravitational potential energy V generated by the sinking of the overburden rock due to the breaking of the roof G .

[0092] The energy V released during the breaking of the tunnel roof is calculated by the following formula:

[0093]

[0094] In the formula,

[0095] V is the energy released during the breaking process of the tunnel roof, J; V G is the gravitational potential energy generated by the sinking of the overburden, J; V ε is the bending strain energy of the cantilever beam at the critical breaking state, N;

[0096] O is the load of the overlying rock stratum on the cantilever beam, N; h s is the tunnel roof sinking height, m; M max is the maximum cantilever beam moment, N·m; l1 is the critical breaking length of the suspended roof in the goaf, m; E is the elastic modulus of the cantilever beam, Pa; I is the moment of inertia of the cantilever beam section, m 4 ; q is the uniformly distributed load of the cantilever beam, N / m.

[0097] S102, based on the energy released during the breaking process of the tunnel roof, the energy release-roof sinking height relationship diagram, the energy release-overburden stratum load relationship diagram, and the energy release-cantilever beam length relationship diagram are drawn respectively, and the above relationship diagrams are analyzed as follows:

[0098] As attached Figure 1 As shown in the figure, with the increase of the rotational sinking height, the energy released by the key overburden layer increases linearly, indicating that the rotational sinking height is a sensitive factor affecting the energy release of the tunnel roof;

[0099] As attached Figure 2 As shown in the figure, with the increase of overburden load, the energy released by the overburden key layer shows a quadratic parabolic growth trend, which is mainly attributed to the overburden load as the main force source, resulting in more bending strain energy accumulated in the cantilever beam structure, and the corresponding increase in the gravitational potential energy generated by the sinking of the overburden;

[0100] As attached Figure 3 As shown in the figure, the increase in the cantilever beam length causes the released energy to grow in the form of a cubic parabola, and the increase rate is significantly accelerated, indicating that the cantilever beam length is the main condition controlling the structural stability and the compressive strength of the top plate.

[0101] In summary, in order to prevent and control rock burst, it is necessary to accelerate the breaking process of the key layer of the roof overburden and reduce the length of the overhanging roof layer. Therefore, cracks need to be set in the key layer of the overburden. Under the action of the deadweight of the rock layer and mining stress, the pre-existing cracks in the cantilever beam will expand, and the expansion and penetration mode directly determines whether the overburden layer will break.

[0102] S2, as attached Figure 4 As shown, the boreholes are arranged at equal intervals along the inclination of the working face, and a cutting angle is formed between the boreholes and the working face. In order to achieve the expected effect of effectively cutting off the suspended rock layer and smoothly collapsing it, and to avoid sudden collapse of the roof due to a large area of ​​suspended roof, which in turn causes disastrous consequences, it is necessary to obtain the minimum cutting angle formed by the boreholes and the working face, and calculate the minimum cutting angle by the following formula:

[0103]

[0104] In the formula,

[0105] a is the length of the pre-existing crack, m; K c is the fracture toughness; λ is the crack extension compression-shear ratio coefficient; π is the circumference of the circumference, which is taken as 3.14; F M(a / b) is the bending force corresponding to the bending moment of the key layer, N; q is the uniformly distributed load of the cantilever beam; l is the length of the basic top lateral cantilever beam, m; θ is the top cutting angle; b1 is the thickness of the key layer, m.

[0106] S3, as attached Figure 7 As shown in FLAC 3DThe numerical model of the non-penetrating pre-cracked top surface of the tunnel roof is established in the software, which specifically includes the following steps:

[0107] S301, construct a numerical model according to the actual engineering dimensions of the mine to be tested. The actual engineering dimensions in this embodiment are 300m in length and 100m in height along the working face. Therefore, the overall dimensions of the numerical model are set to: 500m (X) × 300m (Y) × 100m (Z), where the X direction is the inclination of the working face and the Y direction is the advancement direction of the working face.

[0108] By setting the size of the numerical model according to the actual engineering situation, the influence of the non-penetrating pre-crack cutting top surface on the stability, stress distribution and / or displacement change of the mine roof can be better reflected.

[0109] S302, set the number of overlying rock layers in the numerical model, obtain rock physical and mechanical parameters through geological exploration reports, and calculate the different rock mechanical parameters collected according to their distribution in the actual geological body of the mine to be tested in FLAC 3D Assign values ​​to the unit module (Zone) corresponding to the software.

[0110] In this embodiment, the number of overlying rock layers on the tunnel roof in the numerical model is set to 18 layers, and the coal seam is buried at a depth of 950m; the overlying rock layers are all set to the Mohr-Coulomb elastoplastic material model, and the other geological body mechanical parameters are assigned as shown in the attached figure. Figure 5 shown.

[0111] S303, using FLAC 3D The meshing tool of the software completes the meshing of the numerical model. For numerical models with regular geometric shapes, the mesh size can be made uniform. For numerical models with complex shapes, adaptive or manual meshing can be performed. The numerical model constructed in this embodiment is a regular geometric model, using FLAC 3D The software's built-in division method is 100 (X) × 100 (Y) × 54 (Z), that is, 540,000 units. In this step, for key observation areas, the grid density can be increased to improve the accuracy of images and data during numerical model excavation.

[0112] S304, according to the pre-splitting design scheme in the actual project, the starting point, the end point and the depth of the pre-splitting surface are defined in the numerical model to realize the construction of the non-through pre-splitting top cutting structure.

[0113] Considering the reduction of normal and tangential stiffness of the pre-cracked surface, it can be achieved by 3D The contact surface unit (interface) is set in the software to simulate the mechanical properties of the pre-crack surface.

[0114] In this embodiment, the radius of the crushing zone is 845 mm, and the radius of the fissure zone is 3285 mm, so the damage range of the pre-splitting blasting rock mass is 3.3 m. The strength parameter ratio of the damaged rock mass to the undamaged rock mass is as follows:

[0115] σ c / σ c0 =0.28;σ t / σ t0 =0.41; E / E0=0.625; c / c0=0.54.

[0116] In the formula, σ c is the uniaxial compressive strength of damaged rock mass, Pa; σ c0 is the uniaxial compressive strength of undamaged rock mass, Pa; σ t is the uniaxial tensile strength of the damaged rock mass, Pa; σ t0 is the uniaxial tensile strength of the undamaged rock mass, Pa; E is the elastic modulus of the damaged rock mass, Pa; E0 is the elastic modulus of the undamaged rock mass, Pa; is the friction angle in the damaged rock mass, °; is the friction angle of the undamaged rock mass, °; c is the cohesion of the damaged rock mass, Pa; c0 is the cohesion of the undamaged rock mass, Pa.

[0117] For the target rock layer to be cut by pre-splitting blasting in the numerical model, the mechanical parameters of the rock mass under the blasting damage are shown in the attached figure. Figure 6 shown.

[0118] S305, setting the boundary conditions of the numerical model, specifically: setting a fixed boundary at the bottom of the model, setting its vertical direction, i.e., displacement constraints in the Z direction, to simulate the supporting effect of the earth on the geological body; setting displacement constraints in the horizontal directions, i.e., the X and Y directions, on the side of the model according to actual conditions.

[0119] In this embodiment, displacement constraints are set on the sides and bottom of the model. Since the mining depth is 950m, a vertical uniformly distributed force of 22.5MPa is applied to the upper boundary to represent the gravity load of the overlying rock strata, and a trapezoidal distributed load is applied in the horizontal direction. Since the coal seam is buried deep, the lateral pressure coefficient is selected as 1, and the boundary conditions are that the lower boundary is fixed in the vertical direction, and the front, back, left and right boundaries are fixed in the horizontal direction.

[0120] S306: To eliminate the unbalanced force caused by the initial model construction and parameter assignment and to provide a stable initial state for subsequent dynamic simulations such as excavation or loading, run FLAC 3D The software performs initial equilibrium calculations to allow the model to reach equilibrium under the initial stress state.

[0121] S4, calculating the optimal cutting angle between the drill hole and the working surface in step S2 according to the numerical model established in step S3, specifically comprising the following steps:

[0122] S401, in the non-penetrating pre-splitting top surface numerical model, setting the working face to advance to a certain depth, for example, the advancement depth is 100m;

[0123] S402, setting three groups of drilling top cutting angles in the numerical model to simulate the effect of non-penetrating pre-splitting blasting, specifically: the first group of 50°, 65°, 80° combination angles, the second group of 45°, 60°, 75° combination angles, and the third group of 40°, 55°, 70° combination angles;

[0124] S403, running the numerical model to simulate the top cutting effect of non-penetrating pre-splitting blasting, and analyzing the top cutting angles of three groups of boreholes:

[0125] Attached Figure 8 With attached Figure 9-11 Comparative analysis shows that after the non-penetrating pre-splitting blasting is used to cut the roof, the integrity of the key layer roof is destroyed, so that the energy stored in the coal body is partially released. After the working face is mined, the key layer roof is broken and collapsed in time, shortening the key layer hanging length, thereby cutting off the path for the elastic strain energy to be transmitted from the back of the goaf to the front. At the same time, the pre-splitting blasting releases the elastic strain energy of the key layer roof in front of the coal wall, greatly reducing the degree of energy accumulation in the roof.

[0126] Attached Fig. 9 and attached Fig.10 Analysis shows that when the top cutting angle is set at the first group of 50°, 65°, 80° combination angles and the third group of 40°, 55°, 70° combination angles, it is impossible to completely cut off the energy peak concentration zone of the key layer roof, and the key layer roof energy accumulates on both sides of the non-penetrating pre-splitting blasting top cutting area, which fails to release the key layer roof energy and weaken the impact hazard.

[0127] Attached Fig.11 Analysis shows that when the top cutting angle is set at the second group of 45°, 60°, and 75° combination angles, the top cutting angle penetrates the peak energy band of the key layer roof, allowing the roof energy to be released ahead of the working face, optimizing the temporal and spatial relationship between the roof energy release and the roof collapse, and reducing the risk of roof impact. Therefore, the effect is most significant when the fan-shaped holes are combined with top cutting angles of 45°, 60°, and 75°. At this time, the top cutting angle penetrates the peak energy band of the key layer roof, allowing the roof energy to be released ahead of the working face, optimizing the temporal and spatial relationship between the roof energy release and the roof collapse, reducing the risk of roof impact, and has a significant effect on the release of elastic strain energy of the key layer roof, which can effectively reduce the possibility of rock burst.

[0128] In summary, the best drilling top cutting angle is set as a combination of 45°, 60° and 75° fan-shaped holes. Setting the best drilling top cutting angle can promote the timely and orderly collapse of the roof, effectively alleviate the mining pressure on the working face, and effectively avoid the occurrence of large-scale hanging roof phenomena, as well as technical risks such as casualties and equipment damage caused by the impact of roof collapse.

[0129] S5, according to the optimal drilling top cutting angle formed by the drilling hole and the working surface obtained in step S4, a blasting method is used at the drilling hole to form pre-generated cracks in the key overburden layer, and an advanced non-penetrating pre-splitting top cutting surface is constructed.

[0130] In this embodiment, the siltstone has a uniaxial compressive strength of 45.15 MPa, a tensile strength of 4.89 MPa, and a static Poisson's ratio of 0.3. In order to form non-through pre-existing cracks in the key layer of the overburden without causing through cracks, it is necessary to select the density of the explosives filled in the borehole, the explosive detonation velocity, etc., and accurately calculate the radius of the crushing zone and the radius of the crushing zone generated after the explosives explode. Specifically, the following steps are included:

[0131] S501, blasting will cause rock damage within a certain range, resulting in reduced rock integrity and weakening of rock mechanical parameters. Explosion of explosives mainly forms crushing zone, broken zone, fissure zone and elastic deformation zone in the rock. Pre-splitting blasting and top cutting mainly utilizes the weak surface and fissure formed in the rock by the fissure zone formed by blasting, so the radius of the fissure zone can be approximately regarded as the range of the blasting damage zone.

[0132] Under the condition of uncoupled charge, the strong impact load of the cylindrical charge is:

[0133]

[0134] Right now:

[0135]

[0136] In the formula,

[0137] P is the initial pressure of the shock wave in the rock mass, MPa; P0 is the detonation pressure of the explosive, MPa; K is the charge uncoupling coefficient, which takes a value of 1.5 to 2.5; η is the adiabatic index of the explosion product expansion, which takes a value of 3; l c is the axial coefficient of the charge, which takes a value of 1; n is the pressure increase coefficient when the explosion product expands and collides with the blast hole wall, which takes a value of 10; ρ0 is the density of the explosive, kg / m 3 ;D v is the detonation velocity of the explosive, m / s.

[0138] S502, calculate the dynamic Poisson's ratio of the rock mass within the range of engineering blasting loading rate:

[0139] μd =0.8μ

[0140] In the formula, μ d is the dynamic Poisson's ratio of the rock mass, and μ is the static Poisson's ratio of the rock mass.

[0141] S503, the rock in the blasting and crushing area is mainly subjected to compressive stress, and the rock in the crushing area is mainly subjected to tensile damage. The stress intensity at any point in the rock mass is calculated by the following formula:

[0142]

[0143] In the formula, σ i is the stress intensity at any point in the rock mass; σ r is the radial stress in the rock mass, MPa; b2 is the lateral stress coefficient, μ d is the dynamic Poisson's ratio of the rock mass.

[0144] The radial stress of any particle in the rock mass can be expressed as:

[0145]

[0146] In the formula,

[0147] σ r is the radial stress of any particle in the rock mass; p is the stress of the rock mass before being locally disturbed, Pa; is the ratio distance, r is the distance from the calculation point to the charge center, m; r b is the radius of the blasthole, m; a is the attenuation index of load propagation, The positive and negative signs represent the shock wave area and stress wave area respectively; μ d is the dynamic Poisson's ratio of the rock mass.

[0148] S504, according to the Mises criterion, if the following equation is satisfied, the rock mass is destroyed.

[0149] σ i ≥σ0

[0150]

[0151] Where:

[0152] σ i is the stress intensity at any point in the rock mass; ρ0 is the uniaxial stress failure strength of the rock, MPa; σ cd is the uniaxial compressive strength of rock, MPa; σ td is the uniaxial tensile strength of rock, MPa.

[0153] S505: Under the strong compression of the explosion shock wave, the rocks in the crushing zone are extremely crushed, and cracks appear in the rocks in the crushing zone. Based on steps S501-S504, the crushing zone radius and the crushing zone radius are calculated respectively by the following formulas:

[0154] Crushing zone radius:

[0155]

[0156] Crushing zone radius:

[0157]

[0158] Where:

[0159]

[0160] R c is the radius of the rock mass crushing zone, m; R p is the radius of the rock mass crushing zone, m; ρ0 is the density of the explosive, kg / m 3 ;D v is the detonation velocity of the explosive, m / s; n is the pressure increase coefficient when the explosion product expands and collides with the blast hole wall, which is 10; K is the charge decoupling coefficient, which is 1.5-2.5; l c is the axial coefficient of the charge, which takes the value of 1; B is the comprehensive coefficient related to the lateral stress coefficient of the rock and the dynamic Poisson's ratio; a is the length of the pre-existing crack, m; r b is the blasthole radius, m; σ cd is the uniaxial compressive strength of rock, MPa; σ R is the radial stress on the interface between the crushing zone and the crushing zone, MPa, b2 is the lateral stress coefficient; β is an index related to material crushing characteristics, stress distribution, etc.; μ d is the dynamic Poisson's ratio of the rock mass; σ td is the uniaxial tensile strength of rock, MPa.

[0161] In summary, the main underground waterless blasting operations in my country are currently using coal mine permitted third-grade water gel explosives, and the blasting targets are mainly roof sandstone. The main parameters of coal mine permitted third-grade water gel explosives are: explosive density ρ0 = 1100 ~ 1250kg / m 3 , explosive detonation velocity D v=3200~4200m / s, drug roll diameter 40~63mm. In this embodiment, the uniaxial compressive strength of siltstone is 45.15MPa, the tensile strength is 4.89MPa, and the static Poisson's ratio is 0.3. Substituting it into the formula in this section, the radius of the crushing zone is calculated to be 845mm, and the radius of the fissure zone is 3285mm. Therefore, the damage range of the pre-splitting blasting rock mass is 3.3m. According to the results, the blasting is set to form non-through pre-existing cracks in the key layer of the overburden without generating through cracks.

[0162] S6, pre-existing cracks can be used to control the collapse of the tunnel roof. A reasonable length of pre-existing cracks can ensure that the roof collapses in a predetermined manner and range, and prevent safety accidents such as rock burst caused by sudden collapse of a large area. At the same time, pre-existing cracks can release stress concentration in the rock mass. When the crack length is sufficient, it can effectively reduce the stress level of the rock mass around the mining area.

[0163] The length of the pre-existing cracks formed in the key layer is calculated by the following formula, and the drilling spacing is determined according to the length of the pre-existing cracks. The formula for calculating the length of the pre-existing cracks is as follows:

[0164]

[0165] In the formula,

[0166] K C =λ∑K Ⅰ +|∑K Ⅱ |

[0167]

[0168] P=ρgH

[0169] In the formula,

[0170] a is the length of the pre-existing crack; K C is the fracture toughness; λ is the crack extension compression-shear ratio coefficient; π is the circumference of a circle, which is taken as 3.14; F M(a / b) is the bending force corresponding to the bending moment of the key layer, N; q is the uniform load of the cantilever beam; l is the length of the basic top lateral cantilever beam, m; θ is the pre-crack cutting top angle; b1 is the thickness of the key layer, m; F τ(a / b) is the shear force of the key layer, N; F σ(a / b) is the geometric correction parameter; T is the horizontal extrusion pressure on both sides of the overlying rock layer, N; K Ⅰ is the stress intensity factor at the crack tip of mode I; K Ⅱ is the stress intensity factor at the crack tip of mode II; l' is the periodic pressure step, m; S0 is the length of the working surface, m; P is the overburden load; ρ is the total average density of the overburden, kg / m 3 ; g is the acceleration due to gravity, 9.8m 2 / s; H is the vertical height of the overlying rock layer, m.

[0171] S7, as attached Fig.12 and attached Fig.13 As shown, according to the numerical model of the non-through pre-crack cutting top surface of the tunnel roof established in step S3, the elastic strain energy distribution cloud map of the key layer of the overburden is obtained. The elastic strain energy cloud map includes: the elastic strain energy distribution cloud map of the key layer before the top is cut, and the elastic strain energy distribution cloud map of the key layer of N groups with different drilling top cutting angles.

[0172] According to the elastic strain energy distribution cloud map of the key overburden layer, a relationship diagram between the maximum elastic strain energy and the working face advancement distance is produced.

[0173] According to the relationship diagram between the maximum elastic strain energy and the working face advancement distance, the elastic strain energy of the coal body and the key layer roof of the working face decreases as the non-penetrating pre-splitting top cutting spacing decreases. The optimal pre-splitting top cutting spacing is selected by the following method.

[0174] In this embodiment, in order to study the pressure relief effect of mining stress under different cutting top spacing, three pressure relief schemes are designed, and the cutting top spacing is 8m, 10m, and 12m respectively. Fig.14 As shown in the figure, after comparison: when the spacing is 8m, the coal body energy release efficiency is as high as 25.4%, the key layer roof energy release efficiency is 26.03%, and the peak value of the advanced support pressure is reduced by 12.2%, shortening the influence range of the advanced support pressure by 32%. However, if the non-penetrating pre-splitting top cutting spacing is too small, it is easy to cause a large amount of drilling construction and the roof is easy to break, which is not conducive to the stability of the support system. Combined with the actual project and the economic and efficient non-penetrating pre-splitting blasting top cutting work, the pre-splitting top cutting spacing of 10m was finally selected.

[0175] The following is an actual case, in which inclined blasting of broken top holes was carried out during the mining of the 6310 working face of a coal mine to illustrate the practical application of the method described in the present invention.

[0176] The location of the 6310 working face of a coal mine is as shown in the attached figure. Fig.15 shown.

[0177] The geological parameters of the 6310 working face are as follows:

[0178] Coal seam elevation -905.7~-920.4m;

[0179] Ground elevation +34.1~+40.5m;

[0180] The burial depth is 939.8 to 960.9 m, with an average burial depth of about 950.4 m;

[0181] The working face mainly mines 3 coal, with a coal seam thickness of 7.5 to 10 meters and an average thickness of 9.5 meters;

[0182] The coal seam inclination is 0-4°, with an average of 2°. The measured periodic pressure step distance of the 6310 working face is 23-30m. The thickness of the immediate roof is 29.78m, the average thickness of the siltstone rock layer of the immediate roof key layer is 6.3m, and the fracture toughness K C =1.05MPa·m 1 / 2 , the compression-shear ratio coefficient of crack extension λ=1, and the tunnel width s=5m.

[0183] Calculate the pre-crack cutting angle and pre-crack length of the 6310 working surface:

[0184] The horizontal distance x0=1.2m from the basic top fracture base point to the solid coal seam is calculated by the above step S2;

[0185] The basic top lateral cantilever beam length l=33.4m is calculated through the above step S5;

[0186] Substituting the above data into step S2, it is calculated that θ = 43.68°;

[0187] Ignore the higher-order terms of the collective correction parameters in the calculation of stress intensity factors and take F σ(a / b) =1.12, F τ(a / b) =1.3, F M(a / b) =1.122; the length of the rock beam is l=33.4m; θ is taken as an approximate value of 45°;

[0188] Substituting the above data into step S6, the pre-existing fracture length a=2.12m can be obtained.

[0189] In order to reduce the pressure intensity of the working face, inclined blasting broken top holes are implemented during the 6310 working face mining. At the same time, in order to reduce the stress in the wide coal pillar area of ​​the track section transport level tunnel, inclined broken top holes should be constructed in the wide section coal pillar area, and blasting broken top holes should be arranged in the opening before the 6310 working face mining.

[0190] The specific parameters are as follows:

[0191] The inclination angle of the strike borehole is 75°, and the blasthole diameter is 75mm;

[0192] Spacing between blastholes: The spacing between the inclined top holes of the section transport lane is 15m, with 2 holes arranged in each group, and the spacing between the top holes of the cut-off holes is 10m;

[0193] Blast hole inclination angle: The inclination angle of the deep hole is 75°, and the inclination angle of the shallow hole is 60°. The borehole is perpendicular to the coal wall of the section transport lane and toward the production side.

[0194] The strike borehole has an inclination of 75° and is parallel to the section transport lane and towards the opening eye.

[0195] Hole depth: The depth of the deep hole of the inclined drilling is 30m, and the shallow hole is 26m;

[0196] The depth of the strike borehole is 30m;

[0197] The hole depth of the cut-eye drill is 30m;

[0198] Sealing hole length: The sealing hole length shall not be less than 1 / 3 of the blasting hole depth;

[0199] Charge length: The charge length for deep holes with inclined drilling is 12m, and the charge length for shallow holes is 8m.

[0200] The charge length for the strike borehole is 12m, and the charge length for the cut hole is 12m;

[0201] Charge amount: When using water gel column, the diameter of the water gel column is 63mm, and the charge line density is 3.33kg / m;

[0202] Inclined drilling: The charge for deep holes is 40kg, and the charge for shallow holes is 26.66kg;

[0203] Strike drilling: The charge is 40kg.

[0204] The top-breaking blasting drilling holes are arranged in the transport lanes of the two sections of the working face. The spacing between the strike drilling holes is 8m, with one drilling hole arranged in each group; the spacing between the dip drilling holes is 15m, with two drilling holes arranged in each group; one drilling hole is arranged for the opening, with a spacing of 10m. The drilling plan is shown in the figure below. Fig.16 shown.

[0205] The effect of applying this method to the 6310 working face of the coal mine is verified:

[0206] 1. Drilling hole peek observation blasting before and after drilling TV comparison chart, verify the pressure relief effect. Fig.17 The rock blasting damage zone diagram is shown, and it can be seen that:

[0207] (1) Before blasting, the test borehole wall was smooth and intact, with no cracks. After blasting, the test borehole wall was broken and cracks appeared. There were cracks in the test hole, indicating that the radius of the crack zone c ≥ 3.5m.

[0208] (2) Since the distance between the two blasting holes is 8m, the wall of the test borehole was broken and cracks appeared after blasting. Through the drilling hole, cracks were developed in the test hole, so it can be inferred that the cracks between the two blasting holes have been connected, achieving the expected blasting top breaking effect.

[0209] 2. Use ARAMISM / E microseismic system for monitoring.

[0210] Microseismic sensors are reasonably arranged according to the actual situation on site so that the microseismic monitoring network covers the working face area, so as to better monitor the occurrence of microseismic events in the mining area and its surrounding areas. Fig.18 and Fig.19 This is the evolution curve of the daily total energy and daily frequency of microseismic events in the mining area and surrounding areas during the mining of the 6310 working face.

[0211] Depend on Fig.18 and Fig.19 It can be seen that the total daily energy and frequency of microseismic events show periodic changes as the working face continues to advance. The energy and frequency of microseismic events can better reflect the intensity of the movement of the overlying rock formations. The higher the energy of the microseismic event, the more frequent the vibration, and the more serious the damage to the coal and rock mass.

[0212] according to Fig.18 and Fig.19 From the trend of the middle curve, it is found that during the period when the non-pre-splitting top cutting and pressure relief technology was adopted in the 6310 working face (2022.02.01~2022.06.01), there were 14 pressure comings in both the single-day frequency and the total energy curve. Among them, for the single-day frequency analysis, the maximum cycle pressure coming step distance is 33.8m, the minimum cycle pressure coming step distance is 13m, and the average cycle pressure coming step distance is 22.4m; for the single-day total energy analysis, the maximum cycle pressure coming step distance is 33.8m, the minimum cycle pressure coming step distance is 13m, and the average cycle pressure coming step distance is 22.6m. Through the above analysis, it can be seen that the cycle pressure coming step distance of the 6310 working face is about 22m. According to the detection, during the pressure period of the 6310 working face, the roof of the working face was active, and the energy and frequency increased significantly. After the pressure came, the energy and frequency decreased. The non-penetrating pre-cracking pressure relief technology has a significant effect on the release of elastic strain energy of the key layer roof, and can effectively reduce the possibility of impact ground pressure.

[0213] The above shows and describes the basic principles, main features and advantages of the present invention. Those skilled in the art should understand that the present invention is not limited to the above embodiments, and the above embodiments and descriptions are only for explaining the principles of the present invention. Without departing from the spirit and scope of the present invention, the present invention may have various changes and improvements, and these changes and improvements all fall within the scope of the present invention to be protected.

Claims

1. A method for non-penetrating pre-splitting and top cutting of overburden in a fully mechanized caving working face in a coal mine, characterized in that: The following steps are involved: S1, drill holes are arranged at equal intervals along the inclination of the working surface, and the drill holes form a cutting angle with the working surface; S2, in FLAC 3D The numerical model of the non-penetrating pre-cracked top surface of the tunnel roof is established in the software, which specifically includes the following steps: S201, constructing a numerical model according to the actual engineering dimensions of the mine; S202, set the number of overlying rock layers in the numerical model, set the overlying rock layer to the Mohr-Coulomb elastoplastic material model, and 3D The corresponding unit module of the software assigns the rock physical and mechanical parameters corresponding to the overlying rock layer; S203, using FLAC 3D The software’s meshing tool completes the meshing of the numerical model; S204, according to the pre-splitting design scheme in the actual project, defining the starting point, the end point and the depth of the pre-splitting surface in the numerical model, so as to realize the construction of the non-through pre-splitting top cutting structure; S205, setting boundary conditions of the numerical model, specifically: setting a fixed boundary at the bottom of the numerical model and setting a displacement constraint in the vertical direction; setting a displacement constraint in the horizontal direction on the side of the numerical model according to the actual engineering situation; S206, Run FLAC 3D The software performs initial equilibrium calculations to allow the numerical model to reach equilibrium under the initial stress state; S3, setting N groups of drilling top cutting angles in the numerical model to simulate the effect of non-penetrating pre-splitting blasting, and selecting the required drilling top cutting angles of the drilling hole and the working surface; S4, according to the drilling top cutting angle obtained in step S3, using blasting method at the drilling location to create pre-existing cracks in the overburden key layer, and constructing an advanced non-penetrating pre-crack cutting top surface; S5, calculating the length of pre-existing cracks formed in the key overburden layer, and determining the drilling spacing according to the length of the pre-existing cracks; S6. According to the numerical model of the non-penetrating pre-splitting top cutting surface of the tunnel roof established in step S2, a cloud map of the elastic strain energy distribution of the key layer of the overburden is obtained, including a cloud map of the elastic strain energy distribution of the key layer before the top cutting and a cloud map of the elastic strain energy distribution of the key layer of N groups with different drilling top cutting angles; a relationship diagram between the maximum elastic strain energy and the working face advancement distance is prepared according to the cloud map of the elastic strain energy distribution of the key layer of the overburden, so as to obtain the required pre-splitting top cutting spacing.

2. The method for non-penetrating pre-splitting and top cutting of overburden in fully mechanized caving working face of coal mine according to claim 1, characterized in that: The minimum cutting angle between the drill hole and the working surface is calculated by the following formula: In the formula, a is the length of the pre-existing crack, m; K c is the fracture toughness; λ is the crack extension compression-shear ratio coefficient; π is the circumference of the circumference, which is taken as 3.14; F M(a / b) is the bending force corresponding to the bending moment of the key layer, N; q is the uniformly distributed load of the cantilever beam; l is the length of the basic top lateral cantilever beam, m; θ is the top cutting angle; b1 is the thickness of the key layer, m.

3. The method for non-penetrating pre-splitting and top cutting of overburden in fully mechanized caving working face of coal mine according to claim 1, characterized in that: In step S203, FLAC 3D The numerical model is meshed by the meshing module of the software, or by adaptive meshing, or by manual meshing, or by increasing the mesh density.

4. The method for non-penetrating pre-splitting and top cutting of overburden in fully mechanized caving working face of coal mine according to claim 1, characterized in that: In step S204, by 3D The contact surface unit is set in the software to simulate the mechanical properties of the pre-crack surface; the strength parameter ratio of the damaged rock mass to the undamaged rock mass is set as: c / σ c0 =0.28;σ t / σ t0 =0.41; E / E0=0.625; c / c0=0.54; In the formula, σ c is the uniaxial compressive strength of damaged rock mass, Pa; σ c0 is the uniaxial compressive strength of undamaged rock mass, Pa; σ t is the uniaxial tensile strength of the damaged rock mass, Pa; σ t0 is the uniaxial tensile strength of the undamaged rock mass, Pa; E is the elastic modulus of the damaged rock mass, Pa; E0 is the elastic modulus of the undamaged rock mass, Pa; is the friction angle in the damaged rock mass, °; is the friction angle of the undamaged rock mass, °; c is the cohesion of the damaged rock mass, Pa; c0 is the cohesion of the undamaged rock mass, Pa.

5. The method for non-penetrating pre-splitting and top cutting of overburden in fully mechanized caving working face of coal mine according to claim 1, characterized in that: In step S3, three groups of drilling top cutting angles are set to simulate the non-penetrating pre-splitting blasting effect, specifically: the first group of 50°, 65°, 80° combination angles, the second group of 45°, 60°, 75° combination angles, and the third group of 40°, 55°, 70° combination angles; and key layer elastic strain energy distribution cloud maps are made for the three groups of drilling top cutting angles.

6. A method for non-penetrating pre-splitting and top cutting of overburden in a fully mechanized caving working face in a coal mine according to claim 5, characterized in that: Set 45°, 60° and 75° combined fan-shaped holes as the required cutting angle between the drilling hole and the work surface.

7. The method for non-penetrating pre-splitting and top cutting of overburden in fully mechanized caving working face of coal mine according to claim 1, characterized in that: In step S4, in order to form non-through pre-existing cracks in the key overburden layer without causing through cracks, the radius of the crushing zone and the radius of the crushing zone generated after the explosion of the explosives are calculated by the following steps: S401, calculate the strong impact load of the cylindrical charge under uncoupled charge conditions by the following formula: In the formula, P is the initial pressure of the shock wave in the rock mass, MPa; ρ0 is the density of the explosive, kg / m 3 ;D v is the detonation velocity of the explosive, m / s; K is the charge uncoupling coefficient; η is the adiabatic index of the explosion product expansion; l c is the axial coefficient of the charge; n is the pressure increase coefficient when the explosion product expands and collides with the blast hole wall; S402, calculate the dynamic Poisson's ratio of the rock mass by the following formula: m d =0.8m In the formula, μ d is the dynamic Poisson's ratio of the rock mass, and μ is the static Poisson's ratio of the rock mass; S403, calculate the stress intensity at any point in the rock mass using the following formula: In the formula, σ i is the stress intensity at any point in the rock mass; σ r is the radial stress in the rock mass, MPa; b2 is the lateral stress coefficient, μ d is the dynamic Poisson's ratio of the rock mass; σ r is the radial stress of any particle in the rock mass; p is; is the ratio distance, r is the distance from the calculation point to the charge center, m; r b is the radius of the blasthole, m; a is the attenuation index of load propagation, The positive and negative signs represent the shock wave area and stress wave area respectively; μ d is the dynamic Poisson's ratio of the rock mass; S404, the conditions for rock mass failure are identified by the following formula: s i ≥σ0 Where: σ i is the stress intensity at any point in the rock mass; σ0 is the uniaxial stress failure strength of the rock, MPa; σ cd is the uniaxial compressive strength of rock, MPa; σ td is the uniaxial tensile strength of rock, MPa; S405, based on steps S401-S404, the radius of the crushing zone and the radius of the crushing zone are calculated respectively by the following formulas: Crushing zone radius: Crushing zone radius: Where: R c is the radius of the rock mass crushing zone, m; R p is the radius of the rock mass crushing zone, m; ρ0 is the density of the explosive, kg / m 3 ;D v is the detonation velocity of the explosive, m / s; n is the pressure increase coefficient when the explosion product expands and collides with the blast hole wall; K is the charge decoupling coefficient; l c is the axial coefficient of the charge; B is the comprehensive coefficient related to the lateral stress coefficient of the rock and the dynamic Poisson's ratio; a is the length of the pre-existing crack, m; r b is the blasthole radius, m; σ cd is the uniaxial compressive strength of rock, MPa; σ R is the radial stress on the interface between the crushing zone and the crushing zone, MPa, b2 is the lateral stress coefficient; β is an index related to material crushing characteristics, stress distribution, etc.; μ d is the dynamic Poisson's ratio of the rock mass; σ td is the uniaxial tensile strength of rock, MPa.

8. The method for non-penetrating pre-splitting and top cutting of overburden in fully mechanized caving working face of coal mine according to claim 1, characterized in that: In step S5, the pre-existing crack length is calculated by the following formula: In the formula, K C =λ∑K Ⅰ +|∑K Ⅱ | P=ρgH a is the length of the pre-existing crack; K C is the fracture toughness; λ is the crack extension compression-shear ratio coefficient; π is the circumference; FM (a / b) is the bending force corresponding to the bending moment of the key layer, N; q is the uniformly distributed load of the cantilever beam; l is the length of the basic top lateral cantilever beam, m; θ is the pre-crack cutting angle; b1 is the thickness of the key layer, m; F τ(a / b) is the shear force of the key layer, N; F σ(a / b) is the geometric correction parameter; T is the horizontal extrusion pressure on both sides of the overlying rock layer, N; K Ⅰ is the stress intensity factor at the crack tip of mode I; K Ⅱ is the stress intensity factor at the crack tip of mode II; l' is the periodic pressure step, m; S0 is the working surface length, m; P is the overburden load; ρ is the total average density of the overburden, kg / m 3 ; g is the acceleration due to gravity, 9.8m 2 / s; H is the vertical height of the overlying rock layer, m.

9. The method for non-penetrating pre-splitting and top cutting of overburden in fully mechanized caving working face of coal mine according to claim 1, characterized in that: In step S6, the pre-splitting top cutting distance is set to 10m.

10. A method for non-penetrating pre-splitting and top cutting of overburden in a fully mechanized caving working face in a coal mine according to any one of claims 1 to 9, characterized in that: The feasibility of constructing non-penetrating pre-splitting and cutting of the tunnel roof overburden is verified through the following steps: Step 1: Calculate the energy V released during the breaking of the tunnel roof using the following formula: In the formula, V is the energy released during the breaking process of the tunnel roof, J; V G is the gravitational potential energy generated by the sinking of the overburden, J; V ε is the bending strain energy of the cantilever beam at the critical breaking state, N; Q is the load of the overlying rock on the cantilever beam, N; h s is the tunnel roof sinking height, m; M max is the maximum cantilever beam moment, N·m; l1 is the critical breaking length of the suspended roof in the goaf, m; E is the elastic modulus of the cantilever beam, Pa; I is the moment of inertia of the cantilever beam section, m 4 ; q is the uniformly distributed load of the cantilever beam, N / m. Step 2: Based on the energy released during the breaking of the tunnel roof, the energy release-roof sinking height relationship diagram, the energy release-overburden load relationship diagram, and the energy release-cantilever beam length relationship diagram are drawn respectively, and the above relationship diagrams are concluded: In order to prevent and control impact ground pressure, cracks can be set in the key layer of the overburden, and a non-penetrating pre-crack cutting top of the tunnel roof overburden is constructed.

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