Key layer classification method based on coal mine overlying strata fracture energy release disaster-causing type

Abstract

The invention discloses a key layer classification method based on coal mine overlying strata fracture energy release disaster-causing types, and belongs to the technical field of coal mine dynamic disaster prevention and control, and the method comprises the following steps: 1, calculating overlying strata key layers and fracture step pitches thereof; 2, overlying rock key layer fracture energy release calculation is carried out, an overlying rock fracture energy release theoretical model is established, and the fracture energy release magnitude of each key layer is calculated; and step 3, establishing a coal mine stratum wave velocity field and obtaining an attenuation equation. According to the method, the key layers are divided according to the disaster-causing degree by calculating the magnitude of attenuation of the fracture energy of each key layer of the overlying strata to the energy of the coal seam and combining the energy condition of rock burst occurrence, and the method has guiding significance on prevention and treatment of mine earthquakes and rock burst disasters induced by roof fracture of a coal mine; the roof presplitting parameter of a coal mine can be reasonably optimized in the aspect of roof type rock burst prevention, effective and reasonable allocation of mine resources is facilitated, and waste is avoided.

Description

Technical Field

[0001] The invention belongs to the technical field of coal mine dynamic disaster prevention and control, and in particular relates to a key layer classification method based on coal mine overburden fracture energy release type disaster type. Background Art

[0002] In recent years, with the gradual increase in the intensity of coal mining, dynamic disasters such as rock burst and mine earthquakes induced by the breaking of thick and hard roof in mines have become increasingly prominent. Roof blasting pre-cracking and roof hydraulic fracturing have become the main means of preventing and controlling roof breakage-type dynamic disasters in mines. The existing prevention and control methods are all to pre-cracking the target roof in a centralized manner, and the same pre-cracking method is used for roofs of different layers, lithologies and thicknesses. There is no clear method for distinguishing and selecting the target pre-cracking roof and the pre-cracking degree of the target pre-cracking roof. Due to the different energy release degrees and disaster types of roof fractures, using the same pre-cracking method or pre-cracking parameters for roofs of different layers, lithologies and thicknesses may lead to a huge waste of manpower and material resources.

[0003] The main dynamic disasters induced by roof rupture include rock burst and mine tremor. There is a significant difference in the dynamic manifestation and disaster degree of rock burst and mine tremor. In general, the dynamic manifestation and disaster degree of rock burst are much greater than mine tremor. Therefore, according to the type of disaster induced by the energy release of roof rupture, it is of great significance to classify different types of key overburden layers for quantitative classification and prevention of dynamic disasters in mine roof. Therefore, in order to save mine production costs and guide the rational classification and prevention of key overburden layers in mines, a key layer classification method based on the disaster-inducing type of coal mine overburden rupture was invented.

[0004] Based on this, the present invention designs a key layer classification method based on the disaster type caused by the breaking and release of overburden rock in coal mines to solve the above problems. Summary of the invention

[0005] The purpose of the present invention is to solve the problem that due to the different energy release degrees and disaster types of roof fracture, the use of the same pre-splitting method or pre-splitting parameters for roofs of different strata, lithologies and thicknesses may lead to a huge waste of manpower and material resources, and to propose a key layer classification method based on the disaster type of coal mine overburden fracture energy release.

[0006] In order to achieve the above object, the present invention adopts the following technical solutions:

[0007] A key layer classification method based on the disaster type caused by overburden failure and energy release in coal mines includes the following steps:

[0008] Step 1. Calculation of key overburden layers and their fracture steps;

[0009] Step 2. Calculate the breaking energy release of the key layers of the overburden rock, establish a theoretical model of the breaking energy release of the overburden rock, and calculate the breaking energy release of each key layer;

[0010] Step 3. Establish the coal mine stratum wave velocity field and obtain the attenuation equation, and establish the stratum wave velocity field model based on the mine engineering geological data;

[0011] Step 4. Identify the key layer type.

[0012] As a further description of the above technical solution:

[0013] In the step one, according to the coal and rock occurrence characteristics of the mine, physical parameters such as rock layer thickness, bulk density, elastic modulus and tensile strength are obtained, the key layers of the overburden of the working face are determined, the breaking step of each key layer is calculated, and the length of the working face and the lithology and thickness parameters of each key layer are sorted out and summarized.

[0014] As a further description of the above technical solution:

[0015] The calculation steps of overburden breaking energy in step 2 are as follows:

[0016] ① Calculation of roof accumulated energy:

[0017] After coal seam mining, the roof gradually loses the support of the coal seam below and is exposed. The stress in the stratum is readjusted, and the roof is deformed under the loading of the overlying rock strata. When the internal stress exceeds the tensile strength, fracture instability will occur. The roof is simplified as a transverse force bending roof. The calculation formula of bending strain energy is as follows:

[0018]

[0019] Where: Me is the equivalent bending couple applied by the overlying strata, N·m; θ is the maximum rotation angle of the roof, °; S is the span of the roof, m; E is the elastic modulus of the roof, GPa;

[0020] When the roof is exposed to a certain area, the two ends of the roof break first, and the boundary conditions of the fixed beam are transformed into the boundary conditions of the simply supported beam, and then the middle of the roof breaks; from the bending moment distribution characteristics of the roof under the two models, it can be seen that the bending moment at both ends of the fixed beam is ql 2 / 12, the mid-span bending moment of a simply supported beam is ql 2 / 8, that is, the energy release of the simply supported beam in the middle is the largest; therefore, assuming that there is no relative movement between the overburden and the loaded roof, the loaded roof is simplified to a simply supported beam at both ends with a concentrated force in the middle; under this stress state, the bending strain energy stored before the roof breaks is calculated according to the following formula:

[0021]

[0022] Where: F is the concentrated load on the upper part of the top plate, N;

[0023] The roof plate is subjected to the overburden load and bends and deforms. When the maximum tensile stress inside the beam reaches the tensile strength of the rock layer, the roof plate breaks for the first time. The maximum tensile stress inside the roof plate is:

[0024]

[0025] Where: B is the width of the top plate, m; T is the average thickness of the top plate, m;

[0026] The peak load when the roof fails is:

[0027]

[0028] Where: σb is the tensile strength of rock, MPa;

[0029] Substituting equation (4) into equation (2), the maximum bending strain energy stored in the top plate is:

[0030]

[0031] According to equation (3), the bending strain energy stored before the roof fracture is determined by the lithology (σ b , E) and roof size (B, T, S); in order to conveniently characterize the influence of roof parameters on its own energy storage characteristics, the lithology index Mb and size index Gb of the roof are defined as:

[0032]

[0033] G b =B·T·S(7)

[0034] Then formula (5) can be simplified as:

[0035]

[0036] Where: U εmax is the maximum bending strain energy stored before the specimen fracture, J;

[0037] ② Calculation of energy released when roof breaks:

[0038] A numerical model of the PFC2D three-point bending beam was established; a three-point bending test was conducted, and according to the test results, the material parameters in the numerical simulation were calibrated, and the model was loaded to complete fracture; the rock lithology and size indicators were fitted with the fracture energy release coefficient of the roof sample to obtain the fracture energy release coefficient w of the roof sample;

[0039] Further obtain the energy Ur released by the top plate breaking:

[0040] Ur =w·U e (9)

[0041] Substitute the size index and lithology index parameters of each key layer determined in step 1 into formula (9) to obtain the breaking release energy of each key layer.

[0042] As a further description of the above technical solution:

[0043] The step three specifically includes: first, delineating the area to be simulated on the engineering geological map, collecting the coal and rock layer parameters of the area, including rock layer thickness, rock density, rock elastic modulus and rock Poisson's ratio, and establishing a formation wave velocity field model at a ratio of 1:1; the model boundary is a fully transparent non-reflective boundary, and the discrete grid spacing is dz=2m in the longitudinal direction and dx=2m in the transverse direction; according to the stability condition, the time discrete sampling interval is dt=0.2ms; after the model range is delineated, the formation parameters are converted into longitudinal and transverse wave velocities, and the excitation displacement of the source position is used as the loading input;

[0044] Simulate the energy propagation attenuation process of each key layer, evenly arrange measuring points in the straight line from the earthquake source to the coal wall of the working face, with a spacing of more than 2m between the measuring points, to monitor the energy attenuation coefficient at the measuring point location; draw an attenuation coefficient-transmission distance curve with the earthquake source location as the origin, establish a quantitative relationship between the energy attenuation coefficient of the key layer and the transmission distance, and obtain the energy attenuation coefficient curve transmitted by the rock formation;

[0045] The progressive exponential fitting method is used to fit the energy attenuation law of the key layer, and the theoretical calculation model of formation energy attenuation is obtained. The mathematical form is shown in the following formula:

[0046] γ E = b·e -x / c +a(10)

[0047] Where: x is the distance the shock wave propagates, m;

[0048] Substitute the distance between each key layer and each coal seam into formula (12), calculate the breaking energy attenuation coefficient of each key layer at each coal seam position, and then substitute the breaking energy of each key layer into the following formula:

[0049] U i =E i γ E (11).

[0050] As a further description of the above technical solution:

[0051] In the step 4, rock burst is usually accompanied by mining earthquake, but mining earthquake does not necessarily induce rock burst; therefore, the key layer whose breaking energy reaches the energy condition of inducing rock burst is classified as the key layer that mainly induces mining earthquake; the key layer whose breaking energy is large but does not reach the energy condition of inducing rock burst is classified as the key layer that mainly induces rock burst; the specific classification method is as follows:

[0052] ①Energy conditions for rock burst: When the energy accumulated in the coal in front of the coal wall of the working face and the energy attenuated to the working face by the breaking of the roof are greater than the energy dissipated by the destruction of the coal in front of the coal wall of the working face, rock burst occurs. The criterion for rock burst can be expressed by the following formula:

[0053]

[0054] Where: U c To destroy the elastic deformation energy accumulated in coal, J; U i is the energy of the key layer breaking energy attenuated to the coal seam, J; x is the straight-line distance from the earthquake source to the working face, m; U fmin is the energy dissipated by coal body rupture, J; when K<1, rock burst does not occur; when K≥1, rock burst occurs;

[0055] The energy accumulated in coal can be calculated by the following formula:

[0056]

[0057] Where: k is the stress concentration coefficient; m is the thickness of the coal seam; P0 is the original rock stress; E is the elastic modulus of the coal rock; x0 is the width of the coal body in the plastic zone, x0 = 10m; l is the width of the impact coal body, where l x0, l = x0;

[0058] The energy dissipated by the coal body can be calculated by the following formula:

[0059]

[0060] Where: c is the uniaxial compressive strength of coal rock; α is the energy dissipation coefficient of coal body, K E is the coal rock impact energy index, W ET is the elastic performance index of coal rock;

[0061] ② Classification of key layer types: Substitute the energy of each key layer attenuated to the coal seam into formula (12). When K < 1, the breaking energy of the key layer is not enough to induce rock burst, and this type of key layer is classified as the key layer that mainly induces mine shock; when K ≥ 1, the breaking energy of the key layer meets the conditions for the occurrence of rock burst, and this type of key layer is classified as the key layer that mainly induces rock burst.

[0062] As a further description of the above technical solution:

[0063] When calculating the roof energy storage, the roof span is taken as the roof breaking step, the roof width is taken as the working face dip length, and the average thickness of the roof is taken as the average thickness of the corresponding rock layer in the drilling column diagram;

[0064] The size of the numerical model is 200 mm×40 mm, the span between the lower loading points is 150 mm, and the contact model is a parallel bonding model.

[0065] As a further description of the above technical solution:

[0066] The elastic modulus, tensile strength and compressive strength in the numerical model are consistent with those of the corresponding lithology rocks in indoor tests.

[0067] As a further description of the above technical solution:

[0068] The rock samples used in the indoor test were obtained on site.

[0069] As a further description of the above technical solution:

[0070] The coal and rock layer parameters in the coal mine stratum wave velocity field model, including rock layer thickness, rock density, rock elastic modulus and rock Poisson's ratio, are obtained on site.

[0071] As a further description of the above technical solution:

[0072] The coal mine stratum wave velocity field model is established at a ratio of 1:1, the discrete grid spacing is dz=2m in the longitudinal direction and dx=2m in the transverse direction, and the time discrete sampling interval is dt=0.2ms.

[0073] In summary, due to the adoption of the above technical solution, the beneficial effects of the present invention are:

[0074] 1. In the present invention, by calculating the decay of the breaking energy of each key layer of the overburden to the energy size of the coal seam, combined with the energy conditions for the occurrence of rock burst, the key layers are divided according to the degree of disaster causing, which has guiding significance for preventing and controlling roof breaking-induced mine earthquakes and rock burst disasters in coal mines.

[0075] 2. In the present invention, the key overburden layers are divided based on the degree of disaster-causing energy release from the roof, which is beneficial to the reasonable optimization of roof pre-cracking parameters in the prevention and control of roof-type rock burst in coal mines, and is beneficial to the effective and reasonable allocation of mine resources to avoid waste. BRIEF DESCRIPTION OF THE DRAWINGS

[0076] Figure 1 A flow chart of classification of key layer types of overburden rock based on a key layer classification method of coal mine overburden rock fracture energy release type disaster type proposed by the present invention;

[0077] Figure 2 A schematic diagram of calculation of energy release due to breaking of key layers of overburden rock based on a method for classifying key layers of disaster-causing types of overburden rock due to energy release due to breaking of coal mines proposed by the present invention;

[0078] Figure 3 A rock beam numerical model diagram of a key layer classification method based on the type of disaster caused by overburden failure and energy release proposed by the present invention;

[0079] Figure 4 This is a rock beam numerical parameter calibration diagram for a key layer classification method based on coal mine overburden failure energy release type disaster type proposed by the present invention;

[0080] Figure 5 A schematic diagram of the longitudinal wave velocity field of a coal mine stratum velocity field model based on a key layer classification method for coal mine overburden fracture energy release disaster types proposed by the present invention;

[0081] Figure 6 A schematic diagram of the shear wave velocity field of a coal mine stratum velocity field model based on a key layer classification method for coal mine overburden fracture energy release disaster types proposed by the present invention;

[0082] Figure 7 It is the energy density attenuation coefficient distribution curve of a key layer classification method based on the type of disaster caused by overburden failure and energy release in coal mines proposed by the present invention;

[0083] Figure 8 This is a schematic diagram of the relationship between the energy attenuation coefficient and the transmission distance of a key layer classification method based on the coal mine overburden fracture energy release type disaster type proposed by the present invention. DETAILED DESCRIPTION

[0084] The following is a detailed description of the specific embodiments of the present invention in conjunction with the accompanying drawings, which is a part of this specification and illustrates the principles of the present invention through illustrations. Other aspects, features and advantages of the present invention will become clear from the detailed description. In the drawings referred to, the same or similar parts in different figures are represented by the same reference numerals.

[0085] refer to Figures 1 to 8 As shown, a key layer classification method based on the type of disaster caused by overburden failure and energy release in coal mines in this embodiment is specifically as follows:

[0086] Step 1. Calculation of key overburden layers and their fracture steps

[0087] ① Using the principle of identifying key overburden layers, the key layers above the 2-2 coal seam under the K63 borehole in a coal mine were calculated: key layer 1 is a fine sandstone 12.18 m away from the coal seam and 27.63 m thick; key layer 2 is a medium-grained sandstone 49.67 m away from the coal seam and 56.24 m thick; key layer 3 is a fine sandstone 263.79 m away from the coal seam and 17.28 m thick; key layer 4 is a siltstone 290.92 m away from the coal seam and 17.64 m thick; the main key layer 5 is a sandstone layer group 321.57 m away from the coal seam and 158.32 m thick.

[0088] ② From the test results of rock physical and mechanical indicators in the "Geo-Stress Measurement Report" of the mine, it can be known that the tensile strength of fine sandstone is about 2.5MPa; the thickness is 27.63m, and it is substituted into the following formula to obtain the initial and breaking step distances of each key layer of 2-2 coal (Table 1).

[0089]

[0090] Where: h k is the thickness of the kth hard rock layer, m; σ k is the tensile strength of the kth hard rock layer, MPa; q k is the load borne by the kth hard rock layer, MPa;

[0091] Table 1 Summary of key layer fracture steps

[0092]

[0093] Step 2. Calculation of fracture energy release of key overburden layers

[0094] Substitute the tensile strength and elastic modulus of the roof into formula (6) to calculate the roof rock property index; take the roof span as the roof breaking step, the width of the roof as the working face dip length, and the average thickness of the roof as the average thickness of the corresponding rock layer in the borehole column diagram and substitute them into formula (7) to calculate the roof size index.

[0095] The numerical model of the PFC2D three-point bending beam is established, such as Figure 3 As shown, a three-point bending test was carried out. According to the test results, the material parameters in the numerical simulation were calibrated, and the model was loaded to complete fracture. The rock lithology index was fitted with the fracture energy release coefficient of the roof sample, and the fracture energy release coefficient w of the roof sample was obtained, which can be expressed by the following formula:

[0096]

[0097] The energy released by the roof breaking can be calculated as Ur:U r =w·U e

[0098] The expression of the prediction method of roof fracture energy release is further obtained as follows:

[0099]

[0100] Substituting the size index and lithology index parameters of each key layer determined in step 1 into the above formula, the breaking release energy of each key layer is obtained. The calculation results are listed in Table 2.

[0101] Table 2 Breakage release energy of key overburden layers

[0102]

[0103] Step 3. Establishment of coal mine stratum wave velocity field and attenuation equation

[0104] ① Based on the stratigraphic borehole columnar diagram of the southern wing of a mine, a vibration wave propagation velocity model is established. The model is a two-dimensional plane model with a size of 1000m×610m. The parameters of each rock layer are set according to the actual stratigraphic parameters (borehole K63). After the model range is defined, the stratigraphic parameters are converted into longitudinal and transverse wave velocities, where the longitudinal and transverse wave velocity fields are as follows: Figure 5 and Figure 6 shown.

[0105] ② In the simulation, the center position of each key layer of the source model is selected as the source origin. At each spatial grid point, the kinetic energy density peak value during the elastic wave propagation process is extracted to obtain the distribution curve of the overburden fracture energy density attenuation coefficient, such as Figure 7 shown.

[0106] ③ The progressive exponential fitting method is used to fit the energy attenuation law of the key layer to obtain the theoretical calculation model of formation energy attenuation. The mathematical form is shown in the following formula: γ E =90.27·e -x / 14.58 +2.68

[0107] Substitute the distance between each key layer and each coal seam into formula (12) to calculate the breaking energy attenuation coefficient of each key layer at the position of each coal seam. Then substitute the breaking energy of each key layer into formula (13) to obtain the energy of each key layer attenuated to 2-2 coal, as shown in Table 3.

[0108] Table 3 Calculation of fracture energy attenuation of key overburden layers

[0109]

[0110] Step 4. Identify the key layer type

[0111] ① According to the relevant data of the mine, the vertical stress of the 2-2 coal seam is 17.8MPa, and the thickness of the coal seam is about 8m. The elastic modulus of the coal rock is 3.5GPa, and the uniaxial compressive strength is 26MPa; the average impact energy index KE is 1.78, and the average elastic energy index WET is 10.8.

[0112] Substituting the above parameters and the data in the table into equations (14) to (16), the K value of the fracture energy release of each key layer can be obtained, as shown in Table 4.

[0113] Table 4 K values ​​of key layers of overburden

[0114]

[0115] It can be seen from Table 4 that the K values ​​of the four key layers are all less than 1. Therefore, the above key layers are the key layers that mainly induce mine tremors.

[0116] Of course, the above description is not intended to limit the present invention, and the present invention is not limited to the above examples. It should be noted that, for those skilled in the art, several improvements and changes can be made without departing from the principles of the present invention, and these improvements and changes are also considered to be within the scope of protection of the present invention.

Claims

1. A method for classifying key layers based on the types of disasters caused by overburden failure and energy release in coal mines, characterized in that: The following steps are involved: Step 1. Calculation of key overburden layers and their fracture steps; Step 2. Calculate the breaking energy release of the key layers of the overburden rock, establish a theoretical model of the breaking energy release of the overburden rock, and calculate the breaking energy release of each key layer; Step 3. Establish the coal mine stratum wave velocity field and obtain the attenuation equation, and establish the stratum wave velocity field model based on the mine engineering geological data; Step 4. Identify the key layer type.

2. The key layer classification method based on the disaster type caused by the breaking and releasing energy of overburden in coal mines according to claim 1 is characterized in that: In the step one, according to the coal and rock occurrence characteristics of the mine, the physical parameters of rock layer thickness, bulk density, elastic modulus and tensile strength are obtained, the key layers of overburden of the working face are determined, the breaking step of each key layer is calculated, and the length of the working face and the lithology and thickness parameters of each key layer are sorted out and summarized.

3. The key layer classification method based on the disaster type caused by overburden failure and energy release in coal mines according to claim 1 is characterized in that: The calculation steps of overburden breaking energy in step 2 are as follows: ① Calculation of roof accumulated energy: After coal seam mining, the roof gradually loses the support of the coal seam below and is exposed. The stress in the stratum is readjusted, and the roof is deformed under the loading of the overlying rock strata. When the internal stress exceeds the tensile strength, fracture instability will occur. The roof is simplified as a transverse force bending roof. The calculation formula of bending strain energy is as follows: Where: Me is the equivalent bending couple applied by the overlying strata, N·m; θ is the maximum rotation angle of the roof, S is the span of the roof, m; E is the elastic modulus of the roof, GPa; When the roof is exposed to a certain area, the two ends of the roof break first, and the boundary conditions of the fixed beam are transformed into the boundary conditions of the simply supported beam, and then the middle of the roof breaks; from the bending moment distribution characteristics of the roof under the two models, it can be seen that the bending moment at both ends of the fixed beam is ql 2 / 12, the mid-span bending moment of a simply supported beam is ql 2 / 8, that is, the energy release of the simply supported beam in the middle is the largest; therefore, assuming that there is no relative movement between the overburden and the loaded roof, the loaded roof is simplified to a simply supported beam at both ends with a concentrated force in the middle; under this stress state, the bending strain energy stored before the roof breaks is calculated according to the following formula: Where: F is the concentrated load on the upper part of the top plate, N; The roof plate is subjected to the overburden load and bends and deforms. When the maximum tensile stress inside the beam reaches the tensile strength of the rock layer, the roof plate breaks for the first time. The maximum tensile stress inside the roof plate is: Where: B is the width of the top plate, m; T is the average thickness of the top plate, m; The peak load when the roof fails is: Where: σb is the tensile strength of rock, MPa; Substituting equation (4) into equation (2), the maximum bending strain energy stored in the top plate is: According to equation (3), the bending strain energy stored before the roof fracture is determined by the lithology (σ b , E) and roof size (B, T, S); in order to conveniently characterize the influence of roof parameters on its own energy storage characteristics, the lithology index Mb and size index Gb of the roof are defined as: G b =B·T·S(7) Then formula (5) can be simplified as: Where: U εmax is the maximum bending strain energy stored before the specimen fracture, J; ② Calculation of energy released when roof breaks: A numerical model of the PFC2D three-point bending beam was established; a three-point bending test was conducted, and according to the test results, the material parameters in the numerical simulation were calibrated, and the model was loaded to complete fracture; the rock lithology and size indicators were fitted with the fracture energy release coefficient of the roof sample to obtain the fracture energy release coefficient w of the roof sample; Further obtain the energy Ur released by the top plate breaking: U r =w·U e (9) Substitute the size index and lithology index parameters of each key layer determined in step 1 into formula (9) to obtain the breaking release energy of each key layer.

4. The key layer classification method based on the disaster type caused by the breaking and releasing energy of overburden in coal mines according to claim 1 is characterized in that: The step three specifically includes: first, delineating the area to be simulated on the engineering geological map, collecting the coal and rock layer parameters of the area, including rock layer thickness, rock density, rock elastic modulus and rock Poisson's ratio, and establishing a formation wave velocity field model at a ratio of 1:1; the model boundary is a fully transparent non-reflective boundary, and the discrete grid spacing is dz=2m in the longitudinal direction and dx=2m in the transverse direction; according to the stability condition, the time discrete sampling interval is dt=0.2ms; after the model range is delineated, the formation parameters are converted into longitudinal and transverse wave velocities, and the excitation displacement of the source position is used as the loading input; Simulate the energy propagation attenuation process of each key layer, evenly arrange measuring points in the straight line from the earthquake source to the coal wall of the working face, with a spacing of more than 2m between the measuring points, to monitor the energy attenuation coefficient at the measuring point location; draw an attenuation coefficient-transmission distance curve with the earthquake source location as the origin, establish a quantitative relationship between the energy attenuation coefficient of the key layer and the transmission distance, and obtain the energy attenuation coefficient curve transmitted by the rock formation; The progressive exponential fitting method is used to fit the energy attenuation law of the key layer, and the theoretical calculation model of formation energy attenuation is obtained. The mathematical form is shown in the following formula: c E =b·e -x / c +a(10) Where: x is the distance the shock wave propagates, m; Substitute the distance between each key layer and each coal seam into formula (12), calculate the breaking energy attenuation coefficient of each key layer at each coal seam position, and then substitute the breaking energy of each key layer into the following formula: The i =E i ·γ E (11)。 5. The key layer classification method based on the disaster type caused by the breaking and releasing energy of overburden in coal mines according to claim 1 is characterized in that: In the step 4, rock burst is usually accompanied by mining earthquake, but mining earthquake does not necessarily induce rock burst; therefore, the key layer whose breaking energy reaches the energy condition of inducing rock burst is classified as the key layer that mainly induces mining earthquake; the key layer whose breaking energy is large but does not reach the energy condition of inducing rock burst is classified as the key layer that mainly induces rock burst; the specific classification method is as follows: ①Energy conditions for rock burst: When the energy accumulated in the coal in front of the coal wall of the working face and the energy attenuated to the working face by the breaking of the roof are greater than the energy dissipated by the destruction of the coal in front of the coal wall of the working face, rock burst occurs. The criterion for rock burst can be expressed by the following formula: Where: U c To destroy the elastic deformation energy accumulated in coal, J; U i is the energy of the key layer breaking energy attenuated to the coal seam, J; x is the straight-line distance from the earthquake source to the working face, m; U fmin is the energy dissipated by coal body rupture, J; when K<1, rock burst does not occur; when K≥1, rock burst occurs; The energy accumulated in coal can be calculated by the following formula: Where: k is the stress concentration coefficient; m is the thickness of the coal seam; P0 is the original rock stress; E is the elastic modulus of the coal rock; x0 is the width of the coal body in the plastic zone, x0 = 10m; l is the width of the impact coal body, where l x0, l = x0; The energy dissipated by the coal body can be calculated by the following formula: Where: c is the uniaxial compressive strength of coal rock; α is the energy dissipation coefficient of coal body, K E is the coal rock impact energy index, W ET is the elastic performance index of coal rock; ② Classification of key layer types: Substitute the energy of each key layer attenuated to the coal seam into formula (12). When K < 1, the breaking energy of the key layer is not enough to induce rock burst, and this type of key layer is classified as the key layer that mainly induces mine shock; when K ≥ 1, the breaking energy of the key layer meets the conditions for the occurrence of rock burst, and this type of key layer is classified as the key layer that mainly induces rock burst.

6. The key layer classification method based on the disaster type caused by the breaking energy release of overburden rock in coal mines according to claim 3 is characterized in that: When calculating the roof energy storage, the roof span is taken as the roof breaking step, the roof width is taken as the working face dip length, and the average thickness of the roof is taken as the average thickness of the corresponding rock layer in the drilling column diagram; The size of the numerical model is 200 mm×40 mm, the span between the lower loading points is 150 mm, and the contact model is a parallel bonding model.

7. The key layer classification method based on the disaster type caused by the breaking and releasing energy of overburden rock in coal mines according to claim 6 is characterized in that: The elastic modulus, tensile strength and compressive strength in the numerical model are consistent with those of the corresponding lithology rocks in indoor tests.

8. The key layer classification method based on the disaster type caused by the breaking and releasing energy of overburden rock in coal mines according to claim 7 is characterized in that: The rock samples used in the indoor test were obtained on site.

9. The key layer classification method based on the disaster type caused by the breaking and releasing energy of overburden in coal mines according to claim 4 is characterized in that: The coal and rock layer parameters in the coal mine stratum wave velocity field model, including rock layer thickness, rock density, rock elastic modulus and rock Poisson's ratio, are obtained on site.

10. The key layer classification method based on the disaster type caused by overburden failure energy release in coal mines according to claim 9 is characterized in that: The coal mine stratum wave velocity field model is established at a ratio of 1:1, the discrete grid spacing is dz=2m in the longitudinal direction and dx=2m in the transverse direction, and the time discrete sampling interval is dt=0.2ms.

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