Method for preventing and controlling rock burst through partial filling in full-height goaf of coal seam above 8m

Abstract

The invention belongs to the technical field of mine ground pressure and rock stratum control, and discloses a partial filling rock burst prevention and control method for a full-height goaf of a coal seam above 8 m. The rock stratum position of each hard rock stratum is calculated through a key stratum theory, and the caving height and the limit caving distance of each hard rock stratum are obtained through theoretical calculation; determining the position of a main key layer; and data are introduced to judge whether the key layer conforms to the S-R instability theory or not, the maximum separation space below the main key layer is calculated, and the filling rate and the filling height are determined. According to the method, a goaf part filling method is adopted, the goaf part filling method is coupled with broken and expanded rocks after roof caving, the effect that overlying strata movement is converted from full mining to insufficient mining is achieved, the effect of controlling surface subsidence is achieved by controlling the caving form of a main key layer, large mining height is converted into equivalent small mining height through the method, and the mining efficiency is improved. Filling cost is reduced, and wide application of filling and scour prevention after mining of a coal seam with the thickness of 8 m or above is achieved.

Description

Technical Field

[0001] The present invention belongs to the technical field of mine pressure and stratum control, and in particular relates to a method for preventing and controlling rock burst by partially filling goaf areas in coal seams with a full height of more than 8 m. Background Art

[0002] With the depletion of shallow coal resources, the mining depth of mines in my country has increased year by year, and the problem of mine pressure has become increasingly serious. At present, many coal seams in my country have reached more than 8 meters. When mining at a large depth, due to its large mining height, it is more likely to cause violent movement of the overlying rock strata. When the key layer breaks, the mining area will advance and the supporting pressure on both sides will increase sharply, which is very likely to induce dynamic disasters such as rock burst. Moreover, when the main key layer breaks, it will cause the surface to sink, resulting in damage to the ground environment.

[0003] At present, the main prevention and control method for such disasters is to carry out a full filling operation on the working face. However, in the case of a large mining height on the working face, a full filling operation will lead to huge economic consumption and material waste. However, by partially filling the goaf and collapsing the roof, the characteristics of the collapsed rock fragments and expansion will be utilized, which will greatly reduce the use of filling materials, convert the large mining height of the working face into an equivalent small mining height, fundamentally reduce the concentration of the advance and lateral support pressure of the mining field, reduce or reduce the accumulation of stress and energy, and achieve control of rock burst.

[0004] To sum up, although blindly filling the goaf can control the migration of overburden to a certain extent, the benefits do not outweigh the costs. It seriously overuses the economy and materials, greatly aggravating economic losses.

[0005] Through the above analysis, the problems and defects of the prior art are as follows:

[0006] (1) Currently, backfilling is mainly done at a mining height of 1.3 to 3.5 m. Very few thick coal seams above 3.5 m or even 8 m are backfilled at one time. Some areas do not dare to mine the entire height at one time. Most of them just caving the top coal after mining, which is inefficient and reduces the mining rate. Backfilling operations that almost completely fill the goaf will bring huge economic losses and waste of materials.

[0007] (2) Due to the limitations of on-site conditions and measurement methods, it is impossible to fully reflect the fracture situation of the rock formation on site. Blindly carrying out partial filling is dangerous, and the amount of filling material used is large, which aggravates the use of filling materials. A lot of theoretical analysis, formula calculations and experimental simulation methods are required before the operation. Summary of the invention

[0008] In order to overcome the problems existing in the related art, the disclosed embodiment of the present invention provides a method for preventing and controlling rock burst by partially filling the goaf area of ​​a coal seam above 8m in full height mining, which is particularly suitable for determining the filling height and preventing and controlling rock burst and surface subsidence. It solves the problem of taking partial filling operation when the working face is fully mined at one time, and can effectively suppress the occurrence of surface subsidence and rock burst by utilizing the migration law of overburden and the characteristics of rock crushing and expansion; the technical scheme is as follows:

[0009] The present invention is achieved by: a method for preventing and controlling rock burst by partially filling the goaf area of ​​a coal seam with a full height of more than 8m, the method comprising:

[0010] S1, collect parameters, calculate the rock layer position of each hard rock layer, the collapse height of each hard rock layer and the maximum collapse distance through the key layer theory, and clarify the position of the main key layer;

[0011] S2, set the four-side fixed support conditions, calculate the critical layer breaking distance by Marcus engineering simplified formula, and determine the layer location of the collapse zone and fault zone according to the SR instability judgment theory;

[0012] S3, calculate the maximum separation space below the main key layer, calculate the filling height according to the on-site rock expansion rate, determine the filling rate and perform the filling operation.

[0013] In step S1, the overburden structure is analyzed by the key layer theory to determine the position of the key layer and the location of the main key layer.

[0014] Furthermore, the overburden structure is analyzed through the key layer theory to determine the position of the key layer and the location of the main key layer, including:

[0015] Determine the location of hard rock layers in the overburden that become key layers;

[0016] After determining the location of the hard rock layer, calculate and compare the fracture distances of each rock layer to determine the location of the key layer.

[0017] In step S2, the Marcus engineering simplified formula is used to calculate the breaking distance of the key layer using the four-side fixed support formula to determine the position of the key layer where the breaking occurs.

[0018] In step S2, after determining the limit collapse distance of the key layer, the height of the collapse zone and the fault zone is judged.

[0019] Furthermore, the heights of the collapse zone and the fault zone are judged according to the SR instability theory, wherein the instability conditions include sliding instability conditions and deformation instability conditions.

[0020] Furthermore, the judgment formula of sliding instability condition is:

[0021]

[0022] In the formula, is the friction angle between rock blocks, taking 38°-45°; h is the thickness of the rock block, L is the length of the rock block, a is the ultimate span of the roof along the working face advancing direction;

[0023] The discrimination formula for deformation and instability conditions is:

[0024] Δh>Δ

[0025] In the formula, Δh is the separated layer space height below the key stratum, and Δ is the maximum rotation height allowed for the rotary rock blocks of the masonry beam to maintain stability;

[0026]

[0027] n = σ c / σ t

[0028] In the formula, σ c is the compressive strength, σ t is the tensile strength, and n is the ratio of the compressive strength to the tensile strength, is the ratio of the extrusion strength to the compressive strength; k is the ratio of the rock block extrusion strength to the compressive strength, set to 1 / 2-1 / 3;

[0029]

[0030] In the formula, σ p is the extrusion strength between rock blocks;

[0031]

[0032] In the formula, q is the unit area load borne by the rock block, and α is the fracture angle.

[0033] Furthermore, if key stratum 1 simultaneously meets two instability conditions, the hypothesis does not hold, and key stratum 1 is the caving zone; if key stratum 1 cannot simultaneously meet these two conditions, it belongs to the fracture zone; after calculating and judging to obtain the critical strata positions of the caving zone and the fracture zone, assume that the critical key stratum is i, and the highest key stratum position where fracture occurs is j, where j>1, i≤j≤n, and n is the position of the main key stratum; if j<n, the highest key stratum where fracture occurs is the sub-key stratum, and at this time it is non-full mining; if j = n, the main key stratum fractures, and at this time it is full mining, which will cause bending and sliding of the overlying soft rock and loose layer, resulting in surface subsidence.

[0034] In step S3, after judging the caving zone and the fracture zone by whether the key stratum conforms to the S-R instability theory, assume that key stratum 1 is the basic roof, then the separated layer space height below key stratum 1 is:

[0035] Δh=m-(K-1)∑h

[0036] Where, K is the expansion coefficient of collapsed rock, which is 1.33-1.5; m is the mining height of coal seam, in m; ∑h is the moving layer below the key layer;

[0037] The maximum separation space Δ below the key layer i i The calculation formula is:

[0038] Δ i =H i +m-(H i -h c ) f -h c k c

[0039] In the formula, k f k is the expansion coefficient of the fracture zone rock layer, which is 1.15-1.33; c H is the expansion coefficient of the rock layer in the collapse zone, which is 1.33-1.5; i is the distance between key layer i and coal seam, in m; h c is the height of the collapse zone, in m;

[0040] Assume that the critical key layer is i, and the highest key layer where the fracture occurs is j. Then, before the fracture occurs, the height of the separation space below the highest key layer j is:

[0041] Δh j =m-(K-1)(∑h1+h1+∑h2+h2+…∑h i +h i )-(K l -1)(h i +∑h j )

[0042] In the formula, ∑h i is the cumulative thickness from key layer 1 to key layer i, in m; K l is the expansion coefficient of the fractured rock layer, which is taken as 1.15 to 1.33.

[0043] After the position of the key layer is obtained by the formula, Δh-Δ≤0, m 充填后 Set it as an unknown number and calculate m 充填后 Finally, the filling height C is obtained;

[0044] After simplifying the formula, the filling height C formula is as follows:

[0045] (K-1)(∑h1+h1+∑h2+h2+…∑h i +h i )+(Kl -1)(h i +∑h j )=H j

[0046] In the formula, H j is the height of the collapse zone below the critical layer where the highest failure occurs, in meters;

[0047]

[0048] C 充填高度 =m 原煤层采高 -m 充填后

[0049] The filling rate c is:

[0050]

[0051] In the formula, C 充填高度 is the filling height, m 原煤层采高 Original coal seam height

[0052] In combination with all the above technical solutions, the beneficial effects of the present invention are as follows:

[0053] The method for preventing and controlling rock burst by partially filling the goaf in full-height mining of coal seams above 8m of the present invention first calculates the rock layer position of each hard rock layer through the key layer theory, obtains the collapse height and the limit collapse distance of each hard rock layer through theoretical calculation, and clarifies the layer position of the main key layer; brings in data to determine whether the key layer conforms to the SR instability theory, calculates the maximum separation space below the main key layer, and determines the filling rate and filling height. The present invention adopts a method of partially filling the goaf, so that it couples with the broken and expanded rock after the roof collapses, so as to achieve the effect of changing the movement of the overburden from "full mining" to "non-full mining", controls the collapse morphology of the main key layer to achieve the effect of controlling surface subsidence, and converts the large mining height into an equivalent small mining height through this method; the corresponding filling rate and filling height are obtained through this method, which can realize the wide application of filling and anti-burst after mining of coal seams above 8m. The present invention can help workers understand the underground dynamic disaster forms caused by the overburden structure and the collapse of key layers. It can reduce the concentration of the advance and lateral support pressures of the mining area through partial filling, reduce or reduce the accumulation of stress and energy, and achieve control over rock burst. At the same time, it can reduce the filling cost and achieve control over the migration of overburden, which has important guiding significance for underground safe production and can maximize economic benefits.

[0054] The present invention converts the full mining height of the working face into an equivalent small mining height, converts full mining into incomplete mining, and converts the collapse form into a slow sinking form, thereby increasing the supporting area and supporting force of the goaf space, enhancing the integrity of the roof, thereby inhibiting surface subsidence, reducing the support pressure of the mining field and the retention of high static load energy. This method can effectively control the shear slip of the loose layer of the topsoil, protect surface buildings from damage, prevent surface subsidence and rock burst; reduce the use of filling materials, bring huge economic benefits, and effectively prevent the occurrence of dynamic disasters.

[0055] The present invention can effectively determine the filling rate and filling height when mining the full height of coal seams above 8m at one time, can effectively reduce economic losses, and realize the control of overburden migration and rock burst. The filling materials of the present invention can be effectively reduced in use, and the cost can be saved by about 50% compared with the past. The greatly saved use of filling materials can bring huge economic benefits to mining companies. The present invention reduces the use of filling materials by utilizing the laws of overburden migration and rock collapse characteristics, and only needs to use about 50% of the original filling materials to achieve the effect of controlling surface subsidence. For underground environments, the more filling materials, the better, but the rational use of the laws of overburden migration and rock collapse characteristics can well avoid the waste of filling materials, and can also suppress surface subsidence on the basis of achieving rock burst prevention and control. BRIEF DESCRIPTION OF THE DRAWINGS

[0056] The accompanying drawings herein are incorporated in and constitute a part of the specification, illustrate embodiments consistent with the present disclosure, and together with the description, serve to explain the principles of the present disclosure;

[0057] Figure 1 It is a flow chart of a method for preventing and controlling rock burst by partially filling goaf areas in coal seams with a full height of more than 8m provided by an embodiment of the present invention;

[0058] Figure 2 It is a schematic diagram of a method for preventing and controlling rock burst by partially filling goaf areas in coal seams with a full height of more than 8m provided by an embodiment of the present invention;

[0059] Figure 3 It is a geometrical schematic diagram of the separation space below the key layer in the middle of the span provided by an embodiment of the present invention;

[0060] Figure 4 is a diagram showing the change in the fracture morphology of the key layer before and after filling provided by an embodiment of the present invention;

[0061] Figure 5 is a diagram showing changes in the support pressure in front of the coal wall before and after filling provided by an embodiment of the present invention;

[0062] Figure 6 It is a top plate state and stress transfer diagram before and after the key layer is filled provided by an embodiment of the present invention;

[0063] Figure 7 is a volume schematic diagram of rock after crushing and expansion provided by an embodiment of the present invention;

[0064] Figure 8 is a schematic diagram of surface subsidence before and after the topsoil layer is filled provided by an embodiment of the present invention;

[0065] Fig. 9 is a schematic diagram of numerical simulation modeling provided by an embodiment of the present invention;

[0066] Fig.10 Schematic diagram of subsidence of topsoil layer due to mining at different filling rates provided by an embodiment of the present invention;

[0067] Fig.11 It is a schematic diagram of surface subsidence data with different filling rates provided by an embodiment of the present invention. DETAILED DESCRIPTION

[0068] In order to make the above-mentioned objects, features and advantages of the present invention more obvious and easy to understand, the specific embodiments of the present invention are described in detail below in conjunction with the accompanying drawings. In the following description, many specific details are set forth to facilitate a full understanding of the present invention. However, the present invention can be implemented in many other ways different from those described herein, and those skilled in the art can make similar improvements without violating the connotation of the present invention, so the present invention is not limited by the specific implementation disclosed below.

[0069] The innovation of the present invention is that after calculating the appropriate filling height for filling, the main key layer that was originally fractured is transformed into a slow sinking type, which can inhibit the sinking of the rock layer and loose layer above it. It does not need to spend a lot of money to buy a large amount of filling materials. It only needs to spend about 50% of the original amount to achieve the effect of inhibiting surface subsidence and rock burst prevention and control. The migration law of overburden and the characteristics of rock collapse are fully utilized to achieve the purpose of rock burst prevention and control and reduce the use of filling materials.

[0070] Examples, such as Figure 1 As shown, the method for preventing and controlling rock burst by partially filling the goaf in a coal seam with a full height of more than 8m provided in an embodiment of the present invention comprises the following steps:

[0071] S1, collect parameters, calculate the rock layer position of each hard rock layer, the collapse height of each hard rock layer and the maximum collapse distance through the key layer theory, and clarify the position of the main key layer;

[0072] Analyze the overburden structure through the key layer theory, determine the position of the key layer, and determine the location of the main key layer;

[0073] The first step is to determine the location of the hard rock layer in the overburden that becomes the key layer;

[0074] The second step is to determine the location of the hard rock layer, calculate the fracture distance of each rock layer and compare them to determine the location of the key layer;

[0075] The Marcus engineering simplified formula was further used to calculate the critical layer where fracture would occur, using the four-side fixed support formula, that is, the solid coal fixed support formula on all sides.

[0076] S2, set the four-side fixed support conditions, calculate the critical layer breaking distance by Marcus engineering simplified formula, and determine the layer location of the collapse zone and fault zone according to the SR instability judgment theory;

[0077] The Marcus engineering simplified formula and the four-side fixed support formula are used to calculate the breaking distance of the key layer and determine the position of the key layer where the breaking occurs.

[0078] After determining the ultimate collapse distance of the key layer, the height of the collapse zone and the fault zone is judged.

[0079] The height of the collapse zone and the fault zone is determined according to the SR instability theory, where the instability conditions include sliding instability conditions and deformation instability conditions;

[0080] 1. Sliding and instability conditions;

[0081] 2. Deformation instability conditions.

[0082] Among them, the judgment formula of sliding instability condition is:

[0083]

[0084] In the formula, is the friction angle between the rock blocks, which is 38°-45°; h is the thickness of the rock block, and L is the length of the rock block. a is the limit span of the top plate along the advancing direction of the working face;

[0085] The formula for determining deformation instability conditions is:

[0086] Δh>Δ (2)

[0087] Where Δh is the height of the separation space below the key layer, and Δ is the maximum rotation height allowed for the masonry beam to rotate the rock block to maintain stability;

[0088]

[0089] n=σ c / σ t

[0090] In the formula, σ c is the compressive strength, σ t is the tensile strength, n is the ratio of compressive strength to tensile strength, is the ratio of the extrusion strength to the compressive strength; k is the ratio of the extrusion strength to the compressive strength of the rock mass, and is set to be 1 / 2 to 1 / 3;

[0091]

[0092] In the formula, σ p is the extrusion strength between rock masses;

[0093]

[0094] In the formula, q is the load per unit area borne by the rock mass, and α is the fracture angle.

[0095] S3. Calculate the maximum separation space below the main key stratum, calculate the filling height according to the in-situ rock bulking rate, determine the filling rate and perform the filling operation.

[0096] If key stratum 1 simultaneously satisfies two instability conditions, the hypothesis does not hold, and key stratum 1 is the caving zone; if key stratum 1 cannot simultaneously satisfy these two conditions, it belongs to the fracture zone; after calculating and judging the critical strata positions of the caving zone and the fracture zone, assume that the critical key stratum is i, and the highest key stratum position where fracture occurs is j, where j > 1, i ≤ j ≤ n, and n is the stratum position of the main key stratum; if j < n, the highest key stratum where fracture occurs is the sub-key stratum, and at this time it is non-full mining; if j = n, the main key stratum fractures, and at this time it is full mining, which will cause the bending and sliding of the overlying soft rock and loose layer, resulting in surface subsidence.

[0097] In step S3, after judging the caving zone and the fracture zone by whether the key stratum conforms to the S-R instability theory, assume that key stratum 1 is the basic roof, then the separation space height below key stratum 1 is:

[0098] Δh = m - (K - 1)∑h

[0099] In the formula, K is the bulking coefficient of the caved rock, taking 1.33 - 1.5; m is the coal seam mining height, with the unit of m; Σh is the following layer below the key stratum;

[0100] The maximum separation space Δ i below key stratum i is calculated by the formula:

[0101] Δ i = H i + m - (H i - h c )k f - h c k c

[0102] In the formula, k f is the bulking coefficient of the rock strata in the fracture zone, taking 1.15 - 1.33; kc H is the expansion coefficient of the rock layer in the collapse zone, which is 1.33-1.5; i is the distance between key layer i and coal seam, in m; h c is the height of the collapse zone, in m;

[0103] Assume that the critical key layer is i, and the highest key layer where the fracture occurs is j. Then, before the fracture occurs, the height of the separation space below the highest key layer j is:

[0104] Δh j =m-(K-1)(∑h1+h1+∑h2+h2+…∑h i +h i )-(K l -1)(h i +∑h j )

[0105] In the formula, ∑h i is the cumulative thickness from key layer 1 to key layer i, in m; K l is the expansion coefficient of the fractured rock layer, which is taken as 1.15 to 1.33.

[0106] After the position of the key layer is obtained by the formula, Δh-Δ≤0, m 充填后 Set it as an unknown number and calculate m 充填后 Finally, the filling height C is obtained;

[0107] After simplifying the formula, the filling height C formula is as follows:

[0108] (K-1)(∑h1+h1+∑h2+h2+…∑h i +h i )+(K l -1)(h i +∑h j )=H j

[0109] In the formula, H j is the height of the collapse zone below the critical layer where the highest failure occurs, in meters;

[0110]

[0111] C 充填高度 =m 原煤层采高 -m 充填后

[0112] The filling rate c is:

[0113]

[0114] In the formula, C 充填高度 is the filling height, m原煤层采高 is the original coal seam height.

[0115] The present invention is to convert full mining into insufficient mining, such as Figure 4 As shown, the collapse pattern changes into a slow sinking pattern, increasing the supporting area and supporting force of the mined space.

[0116] like Figure 5 As shown in the figure, the front support pressure of the coal wall is reduced, and the roof integrity is enhanced (rather than destroyed, such as Figure 6 As shown in the figure, filling has changed the characteristics of the original large-scale collapse of the overburden. The equivalent small mining height makes the overburden slowly sink, reducing the elastic energy accumulated in the large-span suspended roof under the condition of large mining height, thereby fundamentally reducing the elastic energy released by the roof breakage and reducing the possibility of rock burst. The arrow in the figure represents the direction of stress transfer. After filling, the stress can be effectively transferred. Before filling, due to the large-span suspended roof, the energy accumulated in it cannot be effectively transferred), thereby inhibiting surface subsidence, reducing the possibility of dynamic disasters such as rock burst, and reducing the support pressure of the mining area and the retention of high static load energy.

[0117] Using the rock expansion characteristics such as Figure 7 Instead of coupling it, by converting a large mining height into an equivalent small mining height, the waste of filling materials can be reduced, the economic benefits can be maximized, and the problem of surface subsidence can be solved.

[0118] like Figure 8 As shown, 1 represents the surface subsidence when no filling operation is performed, 2 and 3 represent the surface subsidence when the filling rate is lower than the calculated filling rate, 4 represents the surface subsidence after the calculated filling rate and the surface subsidence when the filling rate is greater than the filling rate (for example, the calculated filling rate is 40%, if it is increased to 50%, 60%, 70%, 80%, the surface subsidence result is not much different from the filling rate 40% result, or even no difference). This is also verified by numerical simulation. By simulating the typical occurrence conditions of a mining area and comparing them with the results obtained by theoretical calculation, it is found that it is in good agreement with the expected results of the invention, such as Fig. 9 , Fig.10 and Fig.11 shown.

[0119] The invention utilizes the rock crushing and swelling characteristics to couple with the partial filling operation, converting the large mining height into an equivalent small mining height, greatly reducing the waste of filling materials, and making full use of the overburden collapse mechanism to achieve the effect of suppressing surface subsidence. Compared with other filling methods, the present invention can not only control the occurrence of rock burst, making coal mining safer and more efficient, but also achieve the effect of controlling surface subsidence while using the least filling material, thereby protecting surface buildings and roads from damage, greatly promoting the development of mining technology.

[0120] Based on the above technical solution, this aspect uses 3DEC software to conduct simulation experiments. 3DEC software can simulate the migration process of overburden caused by coal mining, and can better show the deformation and damage morphology of the overburden and the dynamic evolution of surface subsidence. The simulation results are as follows: Fig. 9 , 10 , as shown in 11.

[0121] The above description is only a preferred specific implementation manner of the present invention, but the protection scope of the present invention is not limited thereto. Any modifications, equivalent substitutions and improvements made by any technician familiar with the technical field within the technical scope disclosed by the present invention and within the spirit and principles of the present invention should be covered within the protection scope of the present invention.

Claims

1. A method for preventing and controlling rock burst by partially filling the goaf in coal seams with a full height of more than 8m, characterized in that: The method includes: S1. Collect parameters, calculate the rock layer positions where each hard rock layer is located, the caving heights of each hard rock layer, and the ultimate caving distances through the key stratum theory, and clarify the layer position of the main key stratum; S2. Set the four-side fixed support conditions, calculate the breaking distance of the key stratum through the Marcus engineering simplified formula, and judge the layer positions of the caving zone and the fracture zone according to the S-R instability judgment theory; S3. Calculate the maximum separation space below the main key stratum, calculate the filling height according to the in-situ rock dilatancy rate, determine the filling rate and perform the filling operation.

2. The method for preventing and controlling rock burst by partially filling the goaf of a coal seam with a full height of more than 8 m according to claim 1 is characterized in that: In step S1, analyze the overlying rock structure through the key stratum theory, judge the layer position of the key stratum, and determine the location of the main key stratum.

3. The method for preventing and controlling rock burst by partially filling the goaf of a coal seam with a full height of more than 8 m according to claim 2, characterized in that: Analyze the overlying rock structure through the key stratum theory, judge the layer position of the key stratum, and determine the location of the main key stratum, including: Determine the positions of the hard rock layers that become key strata in the overlying rock; After determining the positions of the hard rock layers, calculate the breaking distances of each rock layer and compare them to determine the location of the key stratum.

4. The method for preventing and controlling rock burst by partially filling the goaf of a coal seam with a full height of more than 8 m according to claim 1 is characterized in that: In step S2, use the Marcus engineering simplified formula and adopt the four-side fixed support formula to calculate the breaking distance of the key stratum and judge the layer position of the key stratum that undergoes breaking.

5. The method for preventing and controlling rock burst by partially filling the goaf in coal seams with a full height of more than 8 m according to claim 1, characterized in that: In step S2, after determining the ultimate caving distance of the key stratum, then judge the heights of the caving zone and the fracture zone.

6. The method for preventing and controlling rock burst by partially filling the goaf of a coal seam with a full height of more than 8 m according to claim 5, characterized in that: Judge the heights of the caving zone and the fracture zone according to the S-R instability theory, where the instability conditions include the sliding instability condition and the deformation instability condition.

7. The method for preventing and controlling rock burst by partially filling the goaf of a coal seam with a full height of more than 8 m according to claim 6 is characterized in that: The discriminant formula for the sliding instability condition is: In the formula, is the friction angle between the rock blocks, which is 38°-45°; h is the thickness of the rock block, and L is the length of the rock block. a is the limit span of the top plate along the advancing direction of the working face; The discriminant formula for the deformation instability condition is: Δh>Δ In the formula, Δh is the height of the separation space below the key stratum, and Δ is the maximum rotation height allowed for the rotated rock block of the masonry beam to remain stable; n=σ c / s t In the formula, σ c is the compressive strength, σ t is the tensile strength, n is the ratio of compressive strength to tensile strength, is the ratio of the extrusion strength to the compressive strength; k is the ratio of the rock extrusion strength to the compressive strength, which is set to 1 / 2 to 1 / 3; In the formula, σ p is the compression strength between rock blocks; In the formula, q is the unit area load borne by the rock block, and α is the fracture angle.

8. The method for preventing and controlling rock burst by partially filling the goaf in coal seams with a full height of more than 8 m according to claim 6, characterized in that: If key stratum 1 simultaneously satisfies the two instability conditions, the assumption is not valid, and key stratum 1 is the caving zone; if key stratum 1 cannot simultaneously satisfy these two conditions, it belongs to the fracture zone; after calculating and judging to obtain the critical layer positions of the caving zone and the fracture zone, let the critical key stratum be i, and the highest key stratum layer position that undergoes breaking be j, where j>1, i≤j≤n, and n is the layer position of the main key stratum; if j<n, the highest key stratum that undergoes breaking is the sub-key stratum, and this is non-full mining at this time; if j = n, the main key stratum undergoes breaking, and this is full mining at this time, which will cause the bending and sliding of the overlying soft rock and loose layer, resulting in surface subsidence.

9. The method for preventing and controlling rock burst by partially filling the goaf of a coal seam with a full height of more than 8 m according to claim 8, characterized in that: In step S3, after judging the caving zone and the fracture zone by whether the key stratum conforms to the S-R instability theory, let key stratum 1 be the basic roof, then the height of the separation space below key stratum 1 is: Δh=m-(K - 1)Σh In the formula, K is the dilatancy coefficient of the caved rock, taking 1.33 - 1.5; m is the coal seam mining height, with the unit of m; Σh is the following layer below the key stratum; The maximum separation space Δ below the key layer i i The calculation formula is: Δ i =H i +m-(H i -h c )k f -h c k c In the formula, k f k is the expansion coefficient of the fracture zone rock layer, which is 1.15-1.33; c H is the expansion coefficient of the rock layer in the collapse zone, which is 1.33-1.5; i is the distance between key layer i and coal seam, in m; h c is the height of the collapse zone, in m; Let the critical key stratum be i, and the highest key stratum that undergoes breaking be j, then before breaking, the height of the separation space below the highest key stratum j is: Δh j =m-(K-1)(∑h1+h1+∑h2+h2+…∑h i-1 +h i )-(K l -1)(h i +∑h j ) In the formula, ∑h i is the cumulative thickness from key layer 1 to key layer i, in m; K l is the expansion coefficient of the fractured rock layer, which is taken as 1.15 to 1.

33.

10. The method for preventing and controlling rock burst by partially filling the goaf of a coal seam with a full height of more than 8 m according to claim 9, characterized in that: After the position of the key layer is obtained by the formula, Δh-Δ≤0, m 充填后 Set it as an unknown number and calculate m 充填后 Finally, the filling height C is obtained; After formula simplification, the filling height C formula is obtained as follows: (K-1)(∑h1+h1+∑h2+h2+…∑h i +h i )+(K l -1)(h i +∑h j )=H j In the formula, H j is the height of the collapse zone below the critical layer where the highest failure occurs, in meters; C 充填高度 =m 原煤层采高 -m 充填后 The filling rate c is: In the formula, C 充填高度 is the filling height, m 原煤层采高 is the original coal seam height.

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