Near-field control and far-field isolation methods for regional prevention and control of coal and rock dynamic disasters

Abstract

This invention belongs to the field of coal mining technology, specifically disclosing a method for near-field control and far-field isolation in the prevention and control of coal and rock dynamic disasters. First, geological data of the mining area is collected to determine the key overburden strata. Second, based on the height between the key strata and the coal seam, the length of the longwall working face's diagonal mining range is determined, and the working face is divided. Next, the first longwall working face is mined, and near-field control methods are used to regulate the near-field overburden spatial structure. These near-field control methods include roof cutting for pressure relief, coal seam softening, and floor cutting for pressure relief. During the layout of the first longwall working face, roadway mining faces are used to isolate the far-field overburden spatial structure. Finally, the remaining longwall working faces are subjected to near-field control and far-field isolation in the same manner as above until the dynamic disaster prevention and control of the entire mining area is completed. This invention can reduce the possibility of rockbursts, improve the safety of working face operations, and is applicable to disaster prevention and control in coal mining.

Description

Technical Field

[0001] This invention belongs to the field of coal mining technology and relates to a method for near-field control and far-field isolation of coal and rock dynamic disaster areas. Background Technology

[0002] As coal mining depth increases, the mechanical behavior of the surrounding rock in the mining area becomes more complex, and coal and rock dynamic disasters become increasingly serious. Rockburst, as the most representative typical disaster among coal and rock dynamic disasters, seriously restricts the safe and efficient development of mines.

[0003] As the carrier for the large-scale transfer of the self-weight load of the overlying strata, the overlying rock structure plays a dominant role in the incubation and generation of rockburst disasters in mines. When the overlying strata are ideally soft and loose, their failure process is generally plastic failure, and the evolution of the overlying rock structure and the mining-induced stress under its control during mining exhibits a quasi-static asymptotic evolution process. However, when the overlying rock contains hard strata (critical strata), due to their typical brittle nature, the fracture process is brief and violent, often accompanied by dynamic loading and dynamic phenomena. With the increase in mining depth and intensity, the range of strata related to the generation of coal and rock dynamic disasters has exceeded the scope of the working face and the basic roof in general terms, requiring a broader exploration of the relationship between overlying rock structure and rockburst.

[0004] Chinese invention patent CN 104462659 B discloses a method for analyzing the manifestation of rockburst in hard overburden. First, based on the comprehensive columnar section of the mining area and the physical and mechanical parameters of the overburden, the distribution of key layers is determined. Then, the initial and periodic displacements of the key layers and their distance from the working face are calculated. Next, the free space height and the failure of the key layers are determined. Finally, the characteristics of rockburst manifestation in hard overburden are revealed. Most existing research suggests that rockburst is generated by the superposition of static and dynamic loads on the coal and rock mass. Static loads mainly originate from high ground stress and quasi-static support pressure generated during mining. Dynamic loads mainly originate from rock tremors caused by the failure of hard rock strata. In the source area of ​​the rock tremor, the strain energy released after the failure of hard rock strata forms an energy source similar to an "explosion," and the resulting stress waves propagate to the surrounding rock near the mining face and roadways, creating dynamic disturbances to the mining space. In response, numerous methods for controlling rockbursts have been proposed, primarily targeting factors such as weakening the pressure on the hard roof, coal pillar size, and roadway support. For example, Chinese invention patent CN 109098714 B discloses a method for roadway retention along the goaf in a fully mechanized longwall mining face for high-gas, soft, and extremely unstable coal seams. This method employs active load-bearing support to address the inherent characteristics of high-gas, soft, and extremely unstable coal seams and soft roofs.

[0005] However, from the perspective of overburden structure, there are few methods for preventing and controlling rockbursts that ensure the stability of large-scale overburden spatial structures and fundamentally prevent the generation of large-energy dynamic loads. Summary of the Invention

[0006] The purpose of this invention is to provide a method for preventing and controlling coal and rock dynamic disasters in the near field and isolating the far field from the perspective of overburden structure evolution, so as to reduce the possibility of rockburst and improve the safety of working face operations.

[0007] To achieve the above objectives, the technical solution adopted by this invention is as follows:

[0008] A method for preventing and controlling coal and rock dynamic disasters by near-field regulation and far-field isolation includes the following steps performed sequentially:

[0009] S1. Collect geological data of the well field and determine the stratigraphic position of the key overlying strata;

[0010] S2. Based on the height between the key stratum and the coal seam, determine the length of the longwall working face's dipping range and divide the working face accordingly;

[0011] S3. The first longwall working face was mined, and near-field control methods were used to regulate the near-field overburden space structure. Near-field control methods included roof cutting for pressure relief, coal seam softening, and bottom cutting for pressure relief.

[0012] During the layout of the first longwall working face, the mining roadway working face was used to achieve isolation of the far-field overburden space structure;

[0013] S4. Following the same method as step S3, regulate the near-field overburden space structure and isolate the far-field overburden space structure of the remaining longwall working faces until the dynamic disaster prevention and control of the entire mining area is completed.

[0014] As a limitation, the key layer includes the main key layer and the sub-key layer;

[0015] Step S2 is performed in the following order:

[0016] S21, according to formula According to the failure distance L of the main critical layer k Determine the dipping range of the longwall working face based on the distance from the coal seam height H.

[0017] Mode ;

[0018] Among them, H is the height of the main key layer from the coal seam; L k The fault distance of the main critical layer; The angle of fracture of the rock strata on the side of the goaf; L0 represents the side fracture angle of the working face; L0 represents the dipping mining range of the longwall working face.

[0019] S22, according to formula The length of the working face is determined based on the dip range of the longwall working face, combined with economic factors and mining succession factors.

[0020] Mode ;

[0021] in, Number of working faces;

[0022] S23. The mining area is divided into mining area-based or panel-based layouts.

[0023] As a second limitation, step S3, the process of achieving isolation of the far-field overburden space structure includes the following process: mining the roadway working face while mining the first longwall working face;

[0024] Under normal circumstances, the roadway mining face adopts a continuous mining and filling method of "one mining every other day"; when the roof of the roadway mining face is relatively broken, a mining method of "two mining every three days" is adopted.

[0025] As a further limitation, the working face of the roadway is divided in the following way: x mining roadways are arranged between the transport roadway and the return air roadway, numbered from the inside out as 1, 2, 3...x; among them, odd-numbered roadways are named branch roadways, and even-numbered roadways are named coal pillar roadways, and x≥4;

[0026] The mining sequence of the roadway working face is as follows: joint mining from the inside out. After the branch roadway is completed, the opening point is immediately sealed. After the sealing is completed, the mined branch roadway is filled. After the filling material solidifies, the coal pillar roadway is mined from the inside out. After the coal pillar roadway is mined, the outermost coal pillar roadway is taken as the new cut, and the transport roadway and return airway inside the new cut are sealed.

[0027] As a further limitation, during the mining of branch roadways, when the roof of the branch roadway is intact, a continuous retreat mining method is adopted; when the roof of the branch roadway delaminates or breaks, a skip retreat mining method is adopted.

[0028] During the mining of coal pillar roadways, when the roof of the coal pillar roadway is intact, a continuous retreat mining method is adopted; when the roof of the coal pillar roadway delaminates or breaks, a skip retreat mining method is adopted.

[0029] As a limitation on the roadway mining process, the feature is that during the roadway filling process, the filling system consists of two parts: a surface slurry preparation system and an underground filling system. The filling materials are cement, water, and gangue, and the mass ratio of the three is 0.7:1:12, with the gangue particle size being less than 2cm.

[0030] The branch tunnel is filled from the upper tunnel to the lower tunnel, a grout retaining wall is constructed at the lower end of the branch tunnel, and gangue is filled from the upper end of the branch tunnel.

[0031] As a third limitation, the top-cutting and pressure relief shall be carried out in the following order:

[0032] S31, Construction constant resistance anchor cable

[0033] Construct constant resistance large deformation anchor cables according to the designed support parameters to strengthen the support of the roadway roof;

[0034] S32, Advanced Support

[0035] For the first time, the advance support of the roadway should be no less than 50m; if it is a roadway that is reused, the advance support should be no less than 100m.

[0036] S33, Cutting Hole

[0037] During the mining period, according to the designed drilling parameters, pre-splitting and cutting boreholes were drilled in the roof 50m ahead of the working face.

[0038] S34, Pre-splitting blasting

[0039] According to the design and the blasting parameters determined by the on-site blasting test, the directional pre-splitting blasting device for the top plate was used to carry out pre-splitting and slotting blasting in sequence along the working face advancing direction.

[0040] S35, Rear Lane Support

[0041] After the working face is pushed over, according to the designed support parameters, a rock retaining wall and a temporary roof support system are arranged behind the working face. The temporary roof support system includes individual props, decorative beams, collapsible U-shaped steel, and metal mesh.

[0042] S36. After stabilization, the single-unit roof slab is gradually removed as the working face advances, collapsing and compacting. After stabilization, the single-unit support columns and decorative beams in the stable area are gradually removed.

[0043] S37. Adjust the shape of the tunnel to meet the usage requirements, and spray concrete on the crushed stone side;

[0044] S38, Form the next working face roadway.

[0045] As a fourth limitation, in step S3, the method of softening the coal seam is to inject water by drilling holes from the roadway side to the solid coal side during the longwall mining process; the method of bottom pressure relief is to drill holes from the roadway floor to both sides to relieve pressure.

[0046] The present invention, by adopting the above-described technical solution, achieves the following technical advancements compared to existing technologies:

[0047] (1) This invention proposes a near-field control-far-field isolation method for the prevention and control of coal and rock dynamic disasters from the perspective of overburden structure evolution, which greatly reduces the possibility of rockburst and significantly improves the safety of working face operations;

[0048] (2) This invention starts from the overburden space structure and then designs the working face size with a novel idea; it adopts the roadway mining method to replace the coal pillar method to realize the complete recovery of coal resources in the entire mining area (panel area) and the isolation of the far-field overburden space structure, providing a new method for large-scale overburden space structure control; during the continuous mining of longwall working faces, the methods of cutting the top to relieve pressure, injecting water to soften, and cutting the bottom to relieve pressure are used to realize the control of the near-field overburden space structure; this method starts from the perspective of large and small-scale overburden space structure, greatly reducing the dynamic disasters in the mining area, and has wide applicability for the prevention and control of dynamic disasters in deep conditions containing thick and hard rock strata;

[0049] (3) This invention proposes a near-field control-far-field isolation method for the prevention and control of dynamic disasters. In essence, near-field control is to fundamentally eliminate the generation of high static load + high dynamic load, and far-field isolation fundamentally eliminates the possibility of generating a larger range of high static load + high dynamic load caused by the large-area roof collapse of the main key layer due to the increase of the mining dip scale. Thus, the source control of dynamic disasters such as rock bursts can be achieved.

[0050] This invention belongs to the field of coal mining technology and can reduce dynamic disasters in mining areas and improve the safety of working face operations. Attached Figure Description

[0051] The accompanying drawings are provided to further illustrate the invention and form part of the specification. They are used together with the embodiments of the invention to explain the invention and do not constitute a limitation thereof.

[0052] In the attached diagram:

[0053] Figure 1 This is a schematic diagram illustrating the determination of the inclination range of the long-walled working surface according to an embodiment of the present invention;

[0054] Figure 2 This is a schematic diagram of the working face layout in the mining area according to an embodiment of the present invention;

[0055] Figure 3 a is a schematic diagram of the division of the mining face in an embodiment of the present invention;

[0056] Figure 3 b- Figure 3 d is a schematic diagram of the mining process of the branch roadway in the roadway mining face according to an embodiment of the present invention;

[0057] Figure 3 e- Figure 3g is a schematic diagram of the coal pillar roadway mining process in an embodiment of the present invention;

[0058] Figure 4 This is a schematic diagram of the tunnel mining and backfilling system according to an embodiment of the present invention;

[0059] Figure 5 This is a schematic plan view of the tunnel roof cutting and pressure relief arrangement according to an embodiment of the present invention;

[0060] Figure 6 for Figure 5 Sectional view of the A-A' layout. Detailed Implementation

[0061] The preferred embodiments of the present invention will now be described with reference to the accompanying drawings. It should be understood that the preferred embodiments described herein are for illustrative and explanatory purposes only and are not intended to limit the scope of the invention.

[0062] Example: A method for near-field control and far-field isolation in the prevention and control of coal and rock dynamic disaster areas.

[0063] This embodiment includes the following steps performed sequentially.

[0064] S1. Collect geological data of the well field and determine the stratigraphic position of the key overlying strata;

[0065] S2. Based on the height between the key stratum and the coal seam, determine the length of the longwall working face's dipping range and divide the working face accordingly;

[0066] S3. The first longwall working face was mined, and near-field control methods were used to regulate the near-field overburden space structure. Near-field control methods included roof cutting for pressure relief, coal seam softening, and bottom cutting for pressure relief.

[0067] During the layout of the first longwall working face, the mining roadway working face was used to achieve isolation of the far-field overburden space structure;

[0068] S4. Following the same method as step S3, regulate the near-field overburden space structure and isolate the far-field overburden space structure of the remaining longwall working faces until the dynamic disaster prevention and control of the entire mining area is completed.

[0069] In step S1, during the process of collecting geological data for the mining area, hydrogeological and topographical maps of the mining area, coal seam floor elevations, and mechanical parameters of the coal and rock mass within the mining area are collected. Specifically, this includes 3-5 sets of uniaxial compression tests, Brazilian splitting tests, and shear strength tests of the coal and rock mass. The obtained parameters include: uniaxial compressive, tensile, and shear strengths, elastic modulus, Poisson's ratio, internal friction angle, and cohesion.

[0070] The key stratigraphic position is determined according to existing methods. Key stratigraphic positions include the primary key stratigraphic position and sub-key stratigraphic positions. The primary key stratigraphic position refers to the stratum that controls the overall overlying strata, while the sub-key stratigraphic position refers to the strata that controls the underlying strata. Spatially, the primary key stratigraphic position lies above the sub-key stratigraphic position. The specific process is carried out in the following steps:

[0071] S11, Determine the location of the hard rock strata in the overlying strata.

[0072] Assuming the first rock layer is a hard rock layer, and all layers up to the m-th layer deform in coordination with it, while the (m+1)-th layer does not deform in coordination, then the (m+1)-th layer is the second hard rock layer. Since the first to m-th layers deform in coordination, all layers have the same curvature, forming a composite beam. From the composite beam principle, the load acting on the first hard rock layer can be derived as:

[0073] ①

[0074] In the formula, To account for the load formed by the m-th rock layer on the 1st hard rock layer; h i γ i E i Let be the thickness, unit weight, and elastic modulus of the i-th rock layer, respectively, i = 1, 2, ..., m;

[0075] Considering that the load formed by the (m+1)th layer on the first hard rock layer is

[0076] ②

[0077] Since the (m+1)th layer is a hard rock layer, its deflection is less than that of the rock layers below it. Therefore, the rock layers above the (m+1)th layer no longer need the rock layers below them to bear the load they bear. Thus, it is inevitable that...

[0078] ③

[0079] Substituting equations ① and ② into equation ③ and simplifying, we get:

[0080] ④

[0081] Equation ④ is the formula for determining the location of hard rock strata. Specifically, the determination is made by calculating layer by layer upwards from the first rock stratum above the coal seam. and If equation ④ is satisfied, then the calculation continues upwards. At this point, starting from the first rock layer upwards, the m-th rock layer is the first hard rock layer. Starting from the first hard rock layer, the location of the second hard rock layer is determined using the above method, and so on, until the uppermost hard rock layer is determined and designated as the n-th hard rock layer. By identifying the location of the hard rock layers, the location of the hard rock layers in the overburden and the soft rock layer group they control are obtained.

[0082] S12, fracture distance of hard rock layer

[0083] Based on the mechanical model of a fixed beam and the analysis according to the theory of mechanics of materials, the normal stress at any point within the beam is:

[0084] ⑤

[0085] In the formula, M is the bending moment of the section at any point, in kN•m; y is the distance between any point and the neutral axis of the section, in meters; and h is the thickness of the rock beam, in meters.

[0086] Analysis of the fixed beam shows that the maximum bending moment occurs at both ends of the beam, i.e.

[0087] ⑥

[0088] The corresponding maximum tensile stress is

[0089] ⑦

[0090] when When the rock beam fractures, its ultimate span is obtained from equation ⑦.

[0091] ⑧

[0092] In the formula, h k is the thickness (m) of the kth hard rock layer; is the tensile strength (MPa) of the kth hard rock layer; The load (kN / m) borne by the kth hard rock layer 2 );

[0093] From equation ①, we can see that, It can be determined by the following formula

[0094] 9

[0095] In the formula, the subscript k represents the k-th hard rock layer; the subscript j represents the layer number of the soft rock group controlled by the k-th hard rock layer; m k E represents the number of soft rock layers controlled by the k-th hard rock layer; k,j h k,j γ k,jThese represent the elastic modulus, layer thickness, and unit weight of the j-th rock layer in the soft rock layer group controlled by the k-th hard rock layer, respectively.

[0096] When j=0, it represents the mechanical parameters of the hard rock layer; for example, E. 1,0 h 1,0 γ 1,0 These are the elastic modulus, thickness, and unit weight of the first hard rock layer, E. 1,1 h 1,1 γ 1,1 These are the elastic modulus, thickness, and unit weight of the first layer of soft rock in the soft layer group controlled by the first layer of hard rock;

[0097] Since the elastic modulus of the topsoil layer can be considered as 0, and assuming the thickness of the topsoil layer is H and the unit weight is γ, the load on the uppermost hard rock layer, i.e., the nth hard rock layer, can be calculated using the following formula.

[0098] ⑩

[0099] S13. Compare the fracture distances of each hard rock layer according to the following principles to determine the location of the key layer:

[0100] A. If the k-th hard rock layer is a critical layer, its fracture distance should be less than the fracture distances of all the hard rock layers above it, i.e., it satisfies the following condition.

[0101]

[0102] B. If the fracture distance of the kth hard rock layer is L k If the load is greater than the fracture distance of the (k+1)th hard rock layer above it, then the load borne by the (k+1)th hard rock layer is added to the kth hard rock layer, and the fracture distance of the kth hard rock layer is recalculated.

[0103] C. Determine the layers by starting from the lowest layer of hard rock and working upwards layer by layer. Whether it is true or not, when The fracture distance of the k-th hard rock layer is recalculated.

[0104] In step S2, compared to the prior art which focuses on advancing along the strike of the longwall working face, this embodiment, based on sufficient mining operation (i.e., considering that the strike length is much greater than the dip length), determines that the main key strata within the dip range of the longwall working face must not be breached, thus achieving the purpose of far-field isolation of the overburden space structure. The length of the longwall working face's dip mining range is designed based on the height between the key strata and the coal seam. The specific process is carried out in the following order:

[0105] S21, such as Figure 1 As shown, according to the formula According to the failure distance L of the main critical layer k Determine the dipping range of the longwall working face based on the distance from the coal seam height H.

[0106] Mode ;

[0107] Among them, H is the height of the main key layer from the coal seam; L k The fault distance of the main critical layer; The angle of fracture of the rock strata on the side of the goaf; L0 represents the side fracture angle of the working face; L0 represents the dipping mining range of the longwall working face.

[0108] S22, according to formula The length of the working face is determined based on the dip range of the longwall working face, combined with economic factors and mining succession factors.

[0109] Mode ;

[0110] in, Number of working faces;

[0111] S23, such as Figure 2 As shown, the mining area is divided using either a mining zone layout or a panel layout.

[0112] Step S3, the process of isolating the far-field overburden space structure includes the following process: mining the roadway working face simultaneously during the mining of the first longwall working face;

[0113] After determining the dipping range of the longwall working face, the roadway working face is simultaneously mined during the mining of the first longwall working face. Under normal circumstances, the roadway working face adopts a "one-every-one" continuous mining and filling method. When the roof of the roadway working face is relatively fractured, a "two-every-three" mining method is adopted. The roadway working face is also mined simultaneously with the first longwall working face, and the working face layout of the mining area is as follows: Figure 2 As shown.

[0114] In step S2, the working face is divided as follows: x mining roadways are arranged between the transport roadway and the return air roadway, numbered 1, 2, 3...x from the inside out; odd-numbered roadways are named branch roadways, and even-numbered roadways are named coal pillar roadways, where x ≥ 4; the working face layout plan is shown below. Figure 3 As shown in (a).

[0115] Roadway specifications: The working face branch roadway has a rectangular cross-section with anchor mesh support for the roof; the rough roadway width is 5m, the rough roadway height is 4m, and the rough cross-section is 20m²; the net roadway width is 5m, the net roadway height is 3.8m, and the net cross-section is 19m². The working face coal pillar roadway also has a rectangular cross-section with anchor mesh support for the roof; the rough roadway width is 5m, the rough roadway height is 4m, and the rough cross-section is 20m²; the net roadway width is 5m, the net roadway height is 3.8m, and the net cross-section is 19m².

[0116] Mining sequence: During the mining of the roadway face, a single roadheader works in conjunction with the mining from the inside out. After the branch roadway is completed, the opening point is immediately sealed. After sealing, the mined branch roadway is backfilled. After the backfill material solidifies, the coal pillar roadway is mined from the inside out. After the coal pillar roadway is mined, the outermost coal pillar roadway is taken as the new cut, and the transport roadway and return airway inside the new cut are sealed. Under normal circumstances, the backfill material solidifies for one week.

[0117] During the mining of branch roadways, when the roof of the branch roadway is intact, a continuous retreat mining method is adopted. Continuous retreat: one branch roadway and an adjacent coal pillar roadway constitute a mining unit. Branch roadway mining at the working face proceeds continuously from the inside out, one roadway at a time. When the roof of the branch roadway delaminates or breaks, a skip-retreat mining method is adopted. Skip-retreat: one branch roadway and an adjacent coal pillar roadway constitute a unit. From the inside out, units 1, 2, and 3 are mined first, units 4 and 5 are left unmined, units 6, 7, and 8 are mined, and units 9 and 10 are left unmined. During branch roadway mining at the working face, after continuously mining 3 branch roadways, 2 branch roadways are left unmined as protective coal pillars. Then, 3 more branch roadways are continuously mined, and 2 more are left unmined, and so on. That is, every 2 branch roadways are continuously mined, resulting in 3 branch roadways, or "mining every two and then three." The coal mining procedure at the branch roadway working face is as follows: Figure 3 As shown in (b)-(d).

[0118] During the mining of coal pillar roadways, when the roof of the working face coal pillar roadway is intact, a continuous retreat mining method is adopted. Continuous retreat means that the mining of coal pillar roadways proceeds uninterruptedly from the inside out. When the roof of the working face coal pillar roadway delaminates or breaks, a skip-retreat mining method is adopted. Skip-retreat: First, mining coal pillar roadways #2, #4, and #6, leaving coal pillar roadways #8 and #10, then mining coal pillar roadways #12, #14, and #16... During the mining of coal pillar roadways, three coal pillar roadways are continuously constructed, followed by two coal pillar roadways left unconstructed as protective coal pillars. This process is repeated, meaning three coal pillar roadways are continuously constructed every two coal pillar roadways, i.e., "three-way skip mining." The coal mining procedure at the working face is as follows: Figure 3 As shown in (e)-(g).

[0119] During the tunnel filling process, the main components of the filling material are cement, water, and gangue, with a mass ratio of 0.7:1:12. Simultaneously, to ensure filling strength, a high-efficiency gangue crushing device is installed at the lower end of the gangue bin feeder to ensure that the gangue particle size is less than 2cm before entering the filling system. Taking branch tunnel filling as an example: during the filling process, the branch tunnel (gangue filling tunnel) is filled from the upper tunnel to the lower tunnel. A grout retaining wall is constructed at the lower end of the branch tunnel (lower section, lower position), and gangue filling begins from the upper end of the branch tunnel (upper section, higher position), achieving the filling of one tunnel every 3 days. The specific filling system is as follows... Figure 4 As shown, to ensure the supply of gangue, two filling routes are set up. The filling system consists of two parts: a surface slurry preparation system and an underground filling system. The surface slurry preparation route is: surface slurry preparation station → conveying pipeline (filling borehole) → first mining face roadway; the gangue supply route is: underground gangue bin or gangue from each mining area → crushing station → gangue belt conveyor → first mining face roadway. Figure 4 The diagram shows a backfilling system for coal pillar mining. The backfilling method for coal pillar mining roadways is the same as that for roadways.

[0120] In step S3, the top cutting and pressure relief are carried out in the following sequence.

[0121] S31, Construction constant resistance anchor cable

[0122] Construct constant resistance, large deformation anchor cables according to the designed support parameters to reinforce the roadway roof support; such as Figure 6 Strengthen the support of the roadway roof; if it is a whole-area plan, constant resistance anchor cable support can also be designed in the tunneling roadway.

[0123] S32, Advanced Support

[0124] like Figure 5 For the first time, the advance support of the roadway should be no less than 50m. If it is a roadway that is reused, the advance support should be no less than 100m, depending on the characteristics of the ore pressure manifestation throughout the entire process of the roadway.

[0125] S33, Cutting Hole

[0126] During the mining period, according to the designed drilling parameters, a special slotting drill rig is used to construct pre-splitting slots in the roof 50m ahead of the working face; if the slot depth is appropriate, a regular drill rig can also be used.

[0127] S34, Pre-splitting blasting

[0128] According to the design and the blasting parameters determined by the on-site blasting test, the directional pre-splitting blasting device for the top plate was used to carry out pre-splitting and slotting blasting in sequence along the working face advancing direction.

[0129] S35, Rear Lane Support

[0130] After the working face is pushed over, according to the designed support parameters, a rock retaining wall and a temporary roof support system are arranged behind the working face. The temporary roof support system includes individual props, decorative beams, collapsible U-shaped steel, and metal mesh.

[0131] S36, Stabilization followed by pullback of single unit

[0132] As the working face advances, the roof slab continuously collapses and gradually compacts and stabilizes. After this, the individual support columns and decorative beams in the stable area are gradually removed.

[0133] S37. Adjust the shape of the tunnel to meet the usage requirements, and spray concrete on the crushed stone side;

[0134] S38, Form the next working face roadway.

[0135] In step S3, during the longwall mining process, water is injected into the coal seam by drilling holes on the side of the roadway towards the solid coal side in order to soften the coal seam by water injection; the method of bottom pressure relief is to drill holes on both sides of the roadway floor to relieve pressure.

[0136] This embodiment uses the 1121 working face of a mine in Shandong Province as an example to illustrate the calculation process of the dipping mining range of a longwall working face, where the mining depth is approximately 1000m. The steps include:

[0137] Step 1: Collect geological data of the working face, including information on coal and rock layer thickness and lithology. Prepare multiple standard coal and rock specimens using on-site core drilling. Conduct disc splitting, uniaxial compression, and shear tests on the specimens in the laboratory. The specific implementation steps are as follows:

[0138] Step 101: Prepare 5 to 8 standard coal and rock specimens by drilling in the field;

[0139] Step 102: Conduct disc splitting test, uniaxial compression test and shear test on 5 to 8 standard coal and rock specimens prepared in step 101. Perform each test 4 to 6 times to obtain the corresponding uniaxial tensile strength. The mechanical parameters, including elastic modulus E, uniaxial compressive strength, Poisson's ratio, internal friction angle, and cohesion, were calculated, and their average values ​​were calculated. The calculated average elastic modulus E and uniaxial tensile strength were then used to determine these parameters. Uniaxial compressive strength, Poisson's ratio, internal friction angle, and cohesion were used as the measured coal and rock mechanical parameters.

[0140] Step 103: Based on the rock physical and mechanical parameters, the height between the main key stratum and the coal seam is determined to be approximately 300m, the fracture distance of the main key stratum is approximately 100m, the forward fracture angle is 55°, and the backward fracture angle is 45°. Therefore, the dipping range of the longwall working face can be obtained as follows: ;

[0141] Step 104: Based on the dip range of the longwall working face, and taking into account economic factors and mining succession, the length of the working face is determined: .

Claims

1. A method for near-field control and far-field isolation of coal and rock dynamic disaster areas, characterized in that, Including the following steps performed sequentially, S1. Collect geological data of the well field and determine the stratigraphic position of the key overlying strata; S2. Based on the height between the key stratum and the coal seam, determine the length of the longwall working face's dipping range and divide the working face accordingly; S3. The first longwall working face was mined, and near-field control methods were used to regulate the near-field overburden space structure. Near-field control methods included roof cutting for pressure relief, coal seam softening, and bottom cutting for pressure relief. During the layout of the first longwall working face, the mining roadway working face was used to achieve isolation of the far-field overburden space structure; S4. In the same manner as step S3, regulate the near-field overburden space structure and isolate the far-field overburden space structure of the remaining longwall working faces until the dynamic disaster prevention and control of the entire mining area is completed. The key layer includes the primary key layer and the sub-key layer; Step S2 is performed in the following order: S21, according to formula According to the failure distance L of the main critical layer k Determine the dipping range of the longwall working face based on the distance from the coal seam height H. Mode ; Among them, H is the height of the main key layer from the coal seam; L k The fault distance of the main critical layer; The angle of fracture of the rock strata on the side of the goaf; L0 represents the side fracture angle of the working face; L0 represents the dipping mining range of the longwall working face. S22, according to formula The length of the working face is determined based on the dip range of the longwall working face, combined with economic factors and mining succession factors. Mode ; in, Number of working faces; S23. The mining area shall be divided into mining area-based or panel-based layouts; Step S3, the process of isolating the far-field overburden space structure includes the following process: mining the roadway working face simultaneously during the mining of the first longwall working face; Under normal circumstances, the roadway mining face adopts a continuous mining and filling method of "one mining every other day"; when the roof of the roadway mining face is relatively broken, a mining method of "two mining every three days" is adopted.

2. The method for near-field regulation and far-field isolation of coal and rock dynamic disaster areas according to claim 1, characterized in that, The working face is divided as follows: x mining roadways are arranged between the transport roadway and the return air roadway, numbered from the inside out as 1, 2, 3...x; among them, odd-numbered roadways are named branch roadways, and even-numbered roadways are named coal pillar roadways, where x≥4; The mining sequence of the roadway working face is as follows: joint mining from the inside out. After the branch roadway is completed, the opening point is immediately sealed. After the sealing is completed, the mined branch roadway is filled. After the filling material solidifies, the coal pillar roadway is mined from the inside out. After the coal pillar roadway is mined, the outermost coal pillar roadway is taken as the new cut, and the transport roadway and return airway inside the new cut are sealed.

3. The method for near-field regulation and far-field isolation of coal and rock dynamic disaster areas according to claim 2, characterized in that, During the mining of branch roadways, when the roof of the branch roadway is intact, a continuous retreat mining method is adopted; when the roof of the branch roadway delaminates or breaks, a skip retreat mining method is adopted. During the mining of coal pillar roadways, when the roof of the coal pillar roadway is intact, a continuous retreat mining method is adopted; when the roof of the coal pillar roadway delaminates or breaks, a skip retreat mining method is adopted.

4. The method for near-field regulation and far-field isolation of coal and rock dynamic disaster areas according to claim 2 or 3, characterized in that, During the tunnel filling process, the filling system consists of two parts: a surface slurry preparation system and an underground filling system. The filling materials are cement, water, and gangue, and the mass ratio of the three is 0.7:1:

12. The gangue particle size is less than 2cm. The branch tunnel is filled from the upper tunnel to the lower tunnel, a grout retaining wall is constructed at the lower end of the branch tunnel, and gangue is filled from the upper end of the branch tunnel.

5. The method for near-field control and far-field isolation of coal and rock dynamic disaster areas according to any one of claims 1-3, characterized in that, The depressurization process for cutting off the top should be carried out in the following steps. S31, Construction constant resistance anchor cable Construct constant resistance large deformation anchor cables according to the designed support parameters to strengthen the support of the roadway roof; S32, Advanced Support For the first time, the advance support of the roadway should be no less than 50m; if it is a roadway that is reused, the advance support should be no less than 100m. S33, Cutting Hole During the mining period, according to the designed drilling parameters, pre-splitting and cutting boreholes were drilled in the roof 50m ahead of the working face. S34, Pre-splitting blasting According to the design and the blasting parameters determined by the on-site blasting test, the directional pre-splitting blasting device for the top plate was used to carry out pre-splitting and slotting blasting in sequence along the working face advancing direction. S35, Rear Lane Support After the working face is pushed over, according to the designed support parameters, a rock retaining wall and a temporary roof support system are arranged behind the working face. The temporary roof support system includes individual props, decorative beams, collapsible U-shaped steel, and metal mesh. S36. After stabilization, the single-unit roof slab is gradually removed as the working face advances, collapsing and compacting. After stabilization, the single-unit support columns and decorative beams in the stable area are gradually removed. S37. Adjust the shape of the tunnel to meet the usage requirements, and spray concrete on the crushed stone side; S38, Form the next working face roadway.

6. The method for near-field regulation and far-field isolation of coal and rock dynamic disaster areas according to claim 4, characterized in that, The depressurization process for cutting off the top should be carried out in the following steps. S31, Construction constant resistance anchor cable Construct constant resistance large deformation anchor cables according to the designed support parameters to strengthen the support of the roadway roof; S32, Advanced Support For the first time, the advance support of the roadway should be no less than 50m; if it is a roadway that is reused, the advance support should be no less than 100m. S33, Cutting Hole During the mining period, according to the designed drilling parameters, pre-splitting and cutting boreholes were drilled in the roof 50m ahead of the working face. S34, Pre-splitting blasting According to the design and the blasting parameters determined by the on-site blasting test, the directional pre-splitting blasting device for the top plate was used to carry out pre-splitting and slotting blasting in sequence along the working face advancing direction. S35, Rear Lane Support After the working face is pushed over, according to the designed support parameters, a rock retaining wall and a temporary roof support system are arranged behind the working face. The temporary roof support system includes individual props, decorative beams, collapsible U-shaped steel, and metal mesh. S36, Stabilization followed by pullback of single unit As the working face advances, the roof slab continuously collapses and gradually compacts and stabilizes. After this, the individual support columns and decorative beams in the stable area are gradually removed. S37. Adjust the shape of the tunnel to meet the usage requirements, and spray concrete on the crushed stone side; S38, Form the next working face roadway.

7. The method for near-field control and far-field isolation of coal and rock dynamic disaster areas according to any one of claims 1-3 and 6, characterized in that, In step S3, the method of softening the coal seam is to inject water by drilling holes from the roadway side to the solid coal side during the longwall mining process; the method of bottom pressure relief is to relieve pressure by drilling holes from the roadway floor to both sides.

8. The method for near-field regulation and far-field isolation of coal and rock dynamic disaster areas according to claim 4, characterized in that, In step S3, the method of softening the coal seam is to inject water by drilling holes from the roadway side to the solid coal side during the longwall mining process; the method of bottom pressure relief is to relieve pressure by drilling holes from the roadway floor to both sides.

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