Prevention and control method and system for caving disaster of coal face roof

Abstract

The invention provides a coal face roof caving disaster prevention and control method and system, and the method comprises the steps: obtaining the geological information of each rock stratum corresponding to a stope face open-off cut through the punching of a peephole; based on the geological information, a plurality of pressure relief drill holes are formed in the corresponding coal pillar side close to the open-off cut, and the hard rock stratum of the top plate is presplit through pressure relief treatment, so that the top plate at the open-off cut is converted into a cantilever structure from a beam-shaped structure; monitoring equipment is arranged, and a plurality of buffer cushion layers are arranged on a bottom plate of the goaf; and judging whether to adjust the buffer cushion layer or not by combining the impact data, collected by the monitoring equipment, of the roof during collapse, the first collapse step pitch of the roof under the cantilever structure and the related parameters of the open-off cut. According to the method, the initial caving step pitch of the roof can be reduced, huge dynamic pressure generated during roof caving can be relieved, roof caving disasters are comprehensively prevented and controlled from multiple angles, and the prevention and control effect on the roof caving disasters is improved.

Description

Technical Field

[0001] The present application relates to the technical field of roof disaster prevention and control, and particularly to a method and system for preventing and controlling roof caving disasters in a coal mining face. Background Art

[0002] As people pay more and more attention to the safety of underground mining, roof disaster prevention and control has become an important link in the safe mining process underground. Among them, when the roof caves in, a great deal of energy will be generated, leading to accidents such as the tipping of hydraulic supports and rock bursts in underground roadways.

[0003] In the related art, generally only measures such as installing hydraulic supports are used to prevent and control periodic roof caving. However, in actual application, the prevention and control measures of the above prevention and control schemes are relatively single and cannot meet the prevention and control requirements for roof caving disasters. For example, there are still problems such as a relatively large initial caving step distance and the inability to protect against the huge dynamic pressure brought by roof caving. Summary of the Invention

[0004] The present application aims to solve at least one of the technical problems in the related art to some extent.

[0005] To this end, the first object of the present application is to propose a method for preventing and controlling roof caving disasters in a coal mining face, which can reduce the initial caving step distance of the roof and relieve the huge dynamic pressure generated during roof caving, comprehensively prevent and control roof caving disasters from multiple angles, improve the prevention and control effect of roof caving disasters, and solve problems such as a relatively large initial caving step distance of hard roofs and huge dynamic pressure generated when the overlying surrounding rock of a relatively thick layer caves in.

[0006] The second object of the present application is to propose a system for preventing and controlling roof caving disasters in a coal mining face.

[0007] The third object of the present application is to propose a non-transitory computer-readable storage medium.

[0008] To achieve the above object, the first aspect of the present application is to propose a method for preventing and controlling roof caving disasters in a coal mining face, the method comprising the following steps: Obtain the geological information of each rock stratum corresponding to the cut-eye of the coal mining face by drilling peepholes; Based on the geological information, drill a plurality of pressure relief holes on the side of the first coal pillar corresponding to the cut-eye, and pre-crack the hard roof stratum through pressure relief treatment to convert the roof at the cut-eye from a beam structure to a cantilever structure; A variety of monitoring devices are arranged at different positions in the air return roadway, the conveyor roadway and the gob area, and multiple groups of buffer cushions composed of buffer and energy-absorbing materials are arranged on the floor of the gob area. Among them, the height of the buffer cushion is determined based on the geological information, and the buffer cushion is used to relieve the dynamic pressure generated when the roof collapses to the floor. Combined with the impact data when the roof collapses collected by the variety of monitoring devices, the first caving step distance of the roof under the cantilever structure and the relevant parameters of the cut-through roadway, determine whether to adjust the buffer cushion.

[0009] Optionally, in an embodiment of the present application, the arranging a variety of monitoring devices at different positions in the air return roadway, the conveyor roadway and the gob area includes: arranging multiple groups of borehole stress gauges at the middle positions of the second coal pillar corresponding to the conveyor roadway and the third coal pillar corresponding to the air return roadway respectively, and setting the standard pressure of the borehole stress gauges; arranging multiple groups of dynamic sensors along the parallel direction of the cut-through roadway, wherein each group of dynamic sensors includes a plurality of dynamic sensors arranged along the vertical direction of the cut-through roadway, and the dynamic sensors are used to monitor the impact energy generated when the roof collapses.

[0010] Optionally, in an embodiment of the present application, the combining the impact data when the roof collapses collected by the variety of monitoring devices, the first caving step distance of the roof under the cantilever structure and the relevant parameters of the cut-through roadway, determine whether to adjust the buffer cushion includes: calculating the expected energy generated when the roof collapses according to the first caving step distance of the roof under the cantilever structure, the relevant parameters of the cut-through roadway and the geological information; calculating the average value of the energy data collected by each group of the dynamic sensors, and comparing the calculated energy average value with the expected energy. In the case where the energy average value is in the first interval, replace the buffer cushion; in the case where the energy average value is in the second interval, compare the actual stress values collected by the multiple groups of borehole stress gauges with the standard pressure, and determine the adjustment method of the buffer cushion according to the comparison result of the force data; in the case where the energy average value is in the third interval, keep the buffer cushion unchanged; wherein, the first interval is greater than the second interval, the second interval is greater than the third interval, and the endpoints of each interval are determined according to the expected energy.

[0011] Optionally, in an embodiment of the present application, the determining the adjustment method of the buffer cushion according to the comparison result of the force data includes: in the case where the actual stress value is in the fourth interval, keep the buffer cushion unchanged; in the case where the actual stress value is in the fifth interval, increase the height of the buffer cushion; in the case where the actual stress value is in the sixth interval, replace the buffer cushion; wherein, the sixth interval is greater than the fifth interval, the fifth interval is greater than the fourth interval, and the endpoints of each interval are determined according to the standard pressure.

[0012] Optionally, in an embodiment of the present application, setting multiple groups of buffer cushions composed of buffer energy-absorbing materials on the floor of the goaf includes: calculating the height of the first caving of the roof based on the geological information; determining the height of the buffer cushion according to the flow parameters of the buffer energy-absorbing materials and the height of the first caving of the roof.

[0013] Optionally, in an embodiment of the present application, the pressure relief treatment includes hydraulic fracturing pressure relief and blasting pressure relief. The pre-fracturing of the hard roof strata through the pressure relief treatment includes: drilling multiple short-hole fracturing drill holes on the side of the first coal pillar corresponding to the cut-eye according to the strata strike information of the immediate roof and the main roof above the cut-eye; drilling multiple short-hole fracturing drill holes on the side of the second coal pillar and the side of the third coal pillar according to the geological information of the roof above the second coal pillar and the third coal pillar.

[0014] Optionally, in an embodiment of the present application, after the pre-fracturing of the hard roof strata through the pressure relief treatment, it further includes: setting a single row of hydraulic props on the side of the first coal pillar corresponding to the cut-eye to support the roof at the cut-eye, wherein the spacing between adjacent two hydraulic props in the single row of hydraulic props is the same as the bolt support spacing.

[0015] Optionally, in an embodiment of the present application, obtaining the geological information of each rock stratum corresponding to the cut-eye of the coal mining face through drilling peepholes includes: drilling a peephole perpendicular to the roof at a preset distance along the driving direction of the cut-eye, wherein the depth of the peephole is greater than twice the length of the roof cable bolt; drawing a rock stratum columnar diagram of the cut-eye according to the peephole information obtained from each peephole.

[0016] To achieve the above object, the second aspect of the present application also proposes a prevention and control system for roof caving disasters in a coal mining face, including the following modules: An acquisition module, configured to obtain the geological information of each rock stratum corresponding to the cut-eye of the coal mining face by drilling peepholes; A pre-fracturing module, configured to, based on the geological information, drill multiple pressure relief drill holes on the side of the first coal pillar corresponding to the cut-eye, and pre-fracture the hard roof strata through pressure relief treatment to convert the roof at the cut-eye from a beam structure to a cantilever structure; A setting module, configured to set a variety of monitoring devices at different positions in the air roadway, the machine roadway and the goaf, and set multiple groups of buffer cushions composed of buffer energy-absorbing materials on the floor of the goaf, wherein the height of the buffer cushion is determined based on the geological information, and the buffer cushion is used to relieve the dynamic pressure generated when the roof collapses to the floor; An adjustment module, configured to combine the impact data when the roof collapses collected by the multiple monitoring devices, the initial caving step distance of the roof under the cantilever structure, and the relevant parameters of the cut-through roadway, and determine whether to adjust the buffer cushion layer.

[0017] To implement the above embodiments, a third aspect embodiment of the present application further provides a non-transitory computer-readable storage medium, on which a computer program is stored. When the computer program is executed by a processor, the prevention and control method for roof caving disasters in the first aspect above is implemented.

[0018] The technical solutions provided by the embodiments of the present application at least bring the following beneficial effects: In the present application, pressure relief holes are drilled obliquely upward at a certain angle on the side of the cut-through roadway of the coal mining face close to the coal pillar. The hard rock layer of the roof is pre-cracked by hydraulic fracturing or blasting pressure relief, and the beam-like structure during the initial mining of the coal mining face is transformed into a cantilever structure. Thereby, the initial caving step distance of the roof can be effectively reduced, the excessive caving area of the roof can be avoided, and the pressure manifestation degree of the working face can be reduced. Moreover, as the coal mining face advances, buffer materials are laid on the floor, and the excessive energy generated when the roof collapses is reduced to an acceptable range through the buffer materials, which can prevent various serious accidents caused by huge impact energy in the roadway and maintain the stability of the roadway. Thus, the present application comprehensively prevents and controls roof caving disasters from multiple perspectives, significantly reduces the impact caused by roof caving disasters, improves the reliability, pertinence, and accuracy of the prevention and control of roof caving disasters in the coal mining face, and is beneficial to ensuring safe and stable mining underground in the mine.

[0019] The additional aspects and advantages of the present invention will be partially given in the following description, partially become obvious from the following description, or be understood through the practice of the present invention. Description of the Drawings

[0020] The above and / or additional aspects and advantages of the present application will become obvious and easy to understand from the following description of the embodiments in conjunction with the drawings, where: Figure 1 It is a flowchart of a prevention and control method for roof caving disasters in a coal mining face proposed by an embodiment of the present application; Figure 2 It is a schematic diagram of the principle of preventing and controlling roof caving disasters in a coal mining face proposed by an embodiment of the present application; Figure 3 It is a schematic diagram of the layout of a short-hole hydraulic fracturing process proposed by an embodiment of the present application; Figure 4 It is a schematic diagram of the layout of a long-hole hydraulic fracturing process proposed by an embodiment of the present application; Figure 5 It is a flowchart of a method for adjusting a buffer cushion layer proposed by an embodiment of the present application; Figure 6Schematic diagram of the scenario after the construction of a short-hole hydraulic fracturing process proposed in an embodiment of the present application; Figure 7 Schematic diagram of a drilling and grooving method proposed in an embodiment of the present application; Figure 8 Schematic diagram of the structure of a prevention and control system for the roof caving disaster in a coal mining face proposed in an embodiment of the present application. Detailed implementation manners

[0021] The embodiments of the present invention will be described in detail below. The examples of the embodiments are shown in the drawings, wherein the same or similar reference numerals denote the same or similar elements or elements having the same or similar functions throughout. The embodiments described by referring to the drawings below are exemplary and are intended to explain the present invention, but should not be construed as limiting the present invention.

[0022] It should be noted that, for the convenience of understanding the implementation principle of the roof caving disaster prevention and control method of the present application, the initial caving step distance of the roof targeted by the present application will be described first. The size of the initial caving step distance is determined by the strength of the roof rock formation, the bedding thickness, and the development degree of the joints and fissures in the roof rock formation, and is a comprehensive index of roof stability. During the coal mining process, as the coal mining face advances forward from the open-off cut, the overhanging step distance of the roof gradually increases. Under the action of its own bearing pressure, many through-going fissures are formed inside the overlying rock formation. When the ultimate span is reached, the broken rock blocks rotate and become unstable, resulting in a sharp subsidence of the working face roof, and phenomena such as a generally increased load on the working face supports and even support crushing occur, posing a serious threat to the safety of the working face. In addition, as the coal mining face advances, the roof span gradually increases, and the abutment pressure borne by the coal pillars on the working face side and behind gradually increases, and phenomena such as deformation and spalling of the coal wall will occur.

[0023] At the open-off cut, the roof combines with the coal mining face and the coal pillar behind to form a beam-like structure. Due to the support of the coal pillar behind at the open-off cut, the initial caving step distance is greater than the periodic caving step distance. Therefore, the prevention and control measures for the periodic caving step distance in related embodiments cannot be applied to the initial caving step distance. Especially when the roof lithology is mainly sandstone, the initial caving step distance will be further increased. The larger the initial caving step distance, the more intense the abutment pressure manifestation on the working face, which affects the coal mining process of the working face.

[0024] Therefore, the present application proposes a prevention and control method and system for the roof caving disaster in a coal mining face, which can reduce the initial caving step distance of the hard roof and relieve the huge dynamic pressure generated when the relatively thick overlying surrounding rock caves, and improve the prevention and control effect of the roof caving disaster.

[0025] The following describes a prevention and control method and system for the roof caving disaster in a coal mining face proposed in an embodiment of the present application with reference to the drawings.

[0026] Figure 1 This is a flowchart of a method for preventing and controlling roof caving disasters in a coal mining face proposed in an embodiment of the present application. As Figure 1 shown, the method includes the following steps: Step S101: Obtain the geological information of each rock stratum corresponding to the cut-through roadway of the coal mining face by drilling peepholes.

[0027] Specifically, the present application first obtains the geological data of each rock stratum around the cut-through roadway of the coal mining face (hereinafter referred to as the cut-through roadway for short in the present application) to provide data support and a basis for subsequent prevention and control of roof caving. In order to comprehensively and accurately grasp the geological information of each rock stratum related to the cut-through roadway, such as the rock stratum horizon, the depth, lithology, and development of each rock stratum, etc., the present application obtains information by drilling peepholes.

[0028] It should be noted that the cut-through roadway is a roadway connecting the intake and return air roadways of the coal mining face. In Figure 2 the example shown, it is the area between the coal mining face and the coal pillar. And the rock strata corresponding to the cut-through roadway are the rock strata related to the cut-through roadway around it, including Figure 2 the rock strata above the cut-through roadway in the cross-section shown, as well as the rock strata above the gateways such as the underground air roadway and the machine roadway.

[0029] In an embodiment of the present application, the geological information of each rock stratum corresponding to the cut-through roadway of the coal mining face is obtained by drilling peepholes, including: drilling a peephole perpendicular to the roof at a preset distance along the driving direction of the cut-through roadway, where the depth of the peephole is greater than twice the length of the roof cable bolt; drawing a lithologic column of the cut-through roadway based on the peephole information obtained from each peephole.

[0030] Specifically, during the driving of the first cut-through roadway, along the coal pillar in the driving direction of the cut-through roadway, a peephole is drilled vertically every certain distance (such as 50 m), and the depth of the peephole should be greater than twice the length of the roof cable bolt (generally not less than 15 m). By setting the hole depth in this way, it is possible to ensure the identification of information such as the thickness of the immediate roof and the main roof above the cut-through roadway. Furthermore, summarize the information collected in each peephole, and draw a lithologic column of the entire cut-through roadway based on the position of the peephole and the corresponding rock stratum information, so as to obtain relevant geological information intuitively and conveniently in the future.

[0031] Step S102: Based on the geological information, drill a plurality of pressure relief holes on the side of the first coal pillar corresponding to the cut-through roadway, and pre-crack the hard roof strata through pressure relief treatment to convert the roof at the cut-through roadway from a beam structure to a cantilever structure.

[0032] Specifically, in order to address the problem of excessive initial caving step distance of the hard roof in the coal mining face, this application drills pressure relief holes obliquely upward at a certain angle on the side of the coal pillar near the coal mining face. For example, hydraulic fracturing holes or blasting holes are drilled, and the hard rock layer of the roof is pre-cracked through various pressure relief processes, such as hydraulic fracturing process or blasting pressure relief process. The bearing performance of the first coal pillar 1 in the cut-through roadway on the roof is relieved by means of artificial pressure relief, so as to convert the beam structure at the initial mining of the coal mining face into a cantilever structure.

[0033] Among them, when drilling the pressure relief holes, according to the geological information above the cut-through roadway obtained in the previous step, such as the strike of the immediate roof and the main roof rock layers above the cut-through roadway, the drilling plan of the pressure relief holes is planned. According to the different pressure relief treatment processes adopted, the drilled pressure relief holes can also be of various types. For example, when using the short-hole hydraulic fracturing process in the hydraulic fracturing process, multiple short-hole fracturing holes are drilled, and when using the blasting pressure relief process, blasting holes are drilled.

[0034] It should be noted that after hydraulic fracturing is adopted in this application, it is ensured that the cracks between the fracturing holes penetrate, so as to convert the roof structure at the cut-through roadway from a beam structure to a cantilever structure. Assuming that the height of the impending roof caving is the same, according to the fixed support formula and the cantilever structure formula, in the case of the beam structure, the initial caving step distance of the roof is calculated by the following formula:

[0035] Among them, l 1 is the initial caving step distance under the beam structure, ω max is the maximum deflection, E is the elastic modulus, I is the moment of inertia, q is the uniform load.

[0036] Under the cantilever structure, the initial caving step distance of the roof is calculated by the following formula:

[0037] Among them, l 2 is the initial caving step distance under the cantilever structure, and the other parameters are the same as those in the above formula.

[0038] By comparing these two formulas, it can be seen that the initial caving step distance of the roof at the cut-through roadway is smaller when it is in the cantilever structure, which is more conducive to the initial caving of the mined roof and the problem of long-distance roof suspension will not occur.

[0039] Therefore, this application pre-cracks the roof through pressure relief treatment to obtain Figure 2The prefabricated crack 4 shown can transform the roof structure at the cutting roadway from a beam structure to a cantilever structure, effectively reducing the first caving interval of the roof.

[0040] In one embodiment of the present application, the pressure relief treatment includes hydraulic fracturing pressure relief and blasting pressure relief. The hydraulic fracturing processes that can be adopted include short-hole hydraulic fracturing processes (for example, in the manner shown in Figure 3 ), and long-hole hydraulic fracturing processes (for example, in the manner shown in Figure 4 ). That is to say, when using the artificial pre-fracturing method to relieve pressure on the side of the first coal pillar 1 corresponding to the cutting roadway, short-hole pressure relief, long-hole pressure relief, blasting pressure relief, and various combinations of the above pressure relief methods can be adopted.

[0041] Among them, when adopting the short-hole hydraulic fracturing process, pre-fracturing the hard roof strata includes: according to the strata strike information of the immediate roof and the main roof above the cutting roadway, drilling multiple short-hole fracturing boreholes on the side of the first coal pillar corresponding to the cutting roadway; according to the geological information of the roof above the second coal pillar and the third coal pillar, drilling multiple short-hole fracturing boreholes on the side close to the second coal pillar and the third coal pillar.

[0042] Specifically, as shown in Figure 3 , first, on the side of the first coal pillar 1 corresponding to the cutting roadway, drill multiple short-hole fracturing boreholes 5, the distance between each borehole is Z, and the distance between the borehole and the coal pillar strike is X.

[0043] Then, in order to avoid an overly large triangular hanging roof on the side of the first coal pillar 1 of the working face and the first coal pillar of the return roadway, in this embodiment, according to the roof peeping situation on the side of the second coal pillar 2 in the machine roadway and the third coal pillar 3 in the air roadway close to the cutting roadway (this information has been obtained in step S101), hydraulic fracturing is respectively used for roof cutting on the side close to the second coal pillar 2 and the third coal pillar 3. The short-hole fracturing boreholes 5 at these two places are arranged in a super-backward manner, and the super-backward arrangement is achieved by adjusting the elevation angle of the short-hole fracturing borehole 5 and the included angle with the coal pillar. Similarly, the distance between each short-hole fracturing borehole 5 is Z, and the construction method of the boreholes at these two places is the same as that on the side of the first coal pillar 1.

[0044] Thus, this embodiment can ensure that the entire area around the cutting roadway is a cantilever structure. As the working face is mined, due to the influence of the self-weight of the strata, when the mining distance reaches the first caving interval of the roof under the above cantilever structure, the cutting roadway and the corresponding gateway area will all cave in, and no excessive mine tremors will be caused. The specific drilling method of the boreholes will be described in subsequent embodiments.

[0045] In one embodiment of the present application, after pre-fracturing the hard roof strata through pressure relief treatment, it further includes: arranging a single row of hydraulic props on the side of the first coal pillar corresponding to the cutting roadway to support the roof at the cutting roadway, where the distance between adjacent two hydraulic props in the single row of hydraulic props is the same as the bolt support spacing.

[0046] Specifically, in this embodiment, after the pressure relief treatment, in order to ensure the stability of the roof surrounding rock above the cut - eye, before installing hydraulic supports inside the cut - eye, single - row hydraulic props are first used to support the roof on the side of the first coal pillar 1, and the spacing between each hydraulic prop is the same as the bolt - support spacing of the cut - eye. For example, the spacing between two adjacent hydraulic props can be 0.8 m to 1.2 m. Thus, this embodiment can ensure that the cut - eye will not become unstable due to the conversion to a cantilever structure and avoid the phenomenon of roof fall and collapse.

[0047] Step S103: Set a variety of monitoring devices at different positions in the air roadway, machine roadway and goaf, and set multiple groups of buffer cushions composed of buffer energy - absorbing materials on the floor of the goaf.

[0048] Among them, the height of the buffer cushion is determined based on geological information, and the buffer cushion is used to relieve the dynamic pressure generated when the roof collapses to the floor.

[0049] Specifically, in order to avoid transferring a large amount of energy to the front of the coal mining face when the roof collapses to the floor, resulting in disasters such as deformation of the coal body in the coal mining roadway and affecting the stability of the roadway, this application also lays buffer materials under the goaf as the coal mining face advances, reduces the energy generated when the roof collapses to an acceptable range through the buffer cushion, and uses monitoring devices for monitoring to ensure that the buffer energy - absorbing materials can achieve the expected buffer function.

[0050] As a possible implementation method, in order to ensure that the buffer energy - absorbing materials filled in the goaf can effectively prevent large shock waves generated by the collapse of relatively thick overlying surrounding rock, the height of the first roof collapse can be calculated first according to the characteristics of the roadway surrounding rock, and then the height of the buffer cushion set can be determined according to the height of the first roof collapse. That is, in an embodiment of this application, multiple groups of buffer cushions composed of buffer energy - absorbing materials are set on the floor of the goaf, including: calculating the height of the first roof collapse based on geological information; determining the height of the buffer cushion according to the flow parameters of the buffer energy - absorbing materials and the height of the first roof collapse.

[0051] Specifically, in this embodiment, the height of the first roof collapse of the roadway is calculated first through the following formula:

[0052] Among them, M is the coal seam thickness, is the collapse height of the diameter - top rock stratum, K p is the residual swelling coefficient of the rock.

[0053] Among them, the accumulated height after collapse is , and the void left in the roadway after collapse can be calculated through the above formula , It can be used as the basis for selecting the thickness of the buffer energy-absorbing material. The values of each parameter such as the coal seam thickness in the above formula are the geological information obtained in step S101.

[0054] In addition, after the working face is mined, as the hydraulic support advances forward, the buffer energy-absorbing material is gradually pumped into the goaf at the rear. Therefore, it is required that the fluidity of the buffer material is weak, showing a viscous or solid state. On the basis of the flow requirement, in order to save materials, the filling height can be set to -0.5 m. Thus, the buffer cushion layer 6 shown in Figure 2 can be obtained.

[0055] Based on the above embodiments, in order to avoid the failure of the buffer energy-absorbing material in actual applications, monitoring equipment can be set up before setting the buffer cushion layer 6. In an embodiment of the present application, a variety of monitoring equipment is set at different positions in the air roadway, the machine roadway and the goaf, including: multiple groups of borehole stress gauges are respectively set at the middle positions of the second coal pillar corresponding to the machine roadway and the third coal pillar corresponding to the air roadway, and the standard pressure of the borehole stress gauges is set; multiple groups of dynamic sensors are set along the parallel direction of the cutting face. Among them, each group of dynamic sensors includes multiple dynamic sensors set along the vertical direction of the cutting face, and the dynamic sensors are used to monitor the impact energy generated when the roof collapses.

[0056] For example, as Figure 3 shown, 3 groups of borehole stress gauges are respectively installed in the air roadway and the machine roadway, and the installation distance between adjacent two groups of borehole stress gauges is W. The borehole stress gauges are placed at the middle positions of the coal pillars where they are located to ensure that the stress is not distorted.

[0057] Then, after the borehole stress gauges are installed, the borehole stress gauges are pre-pressurized, and the average value of the 3 groups of borehole stress gauges is first set to F = H × γ , where H is the roadway burial depth, γ is the unit weight, F is the set standard pressure.

[0058] Furthermore, as Figure 2 shown, multiple dynamic sensors 7 are added between the floor and the set buffer cushion layer 6, and data such as the impact force and impact energy of the rock stratum collapse are monitored through the dynamic sensors 7. Specifically, when arranging, at the cutting face, 3 dynamic sensors 7 are installed in a group in the direction perpendicular to the cutting face, and a group of dynamic sensors is installed at a distance of not less than 50 m in the direction parallel to the cutting face. After taking the average value of the data collected by each group of dynamic sensors, the energy average value is obtained, denoted as Q 1 .

[0059] Step S104: Based on the impact data collected by multiple monitoring devices when the roof collapses, the first caving interval of the roof under the cantilever structure, and the relevant parameters of the cut - eye, determine whether to adjust the buffer cushion.

[0060] Specifically, based on the impact data actually collected by multiple monitoring devices when the roof collapses, the determined first caving interval of the roof under the cantilever structure and geological information, as well as the relevant parameters of the cut - eye, determine whether the current buffer cushion 6 is effective, and then make corresponding adjustments to the buffer cushion 6 according to the judgment result.

[0061] To more clearly illustrate the implementation process of determining whether to adjust the buffer cushion in this application, the following uses an adjustment method of a buffer cushion proposed in an embodiment of this application as an example for illustration. Figure 5 The flowchart of an adjustment method of a buffer cushion proposed in an embodiment of this application is as Figure 5 shown, and this method includes the following steps: Step S501: Calculate the expected energy generated when the roof collapses based on the first caving interval of the roof under the cantilever structure, the relevant parameters of the cut - eye, and geological information.

[0062] Specifically, based on the determined first caving interval of the roof under the cantilever structure l 2 , the width of the cut - eye l , the widths of the two gate roads (i.e., the air roadway and the machine roadway) on both sides of the cut - eye l 3 , the caving height of the rock stratum and the coal seam thickness M , calculate the expected energy generated when the roof collapses through the following formula: Q = l 2 × × M ×( l + 2 l 3 )× g where g is the acceleration of free fall.

[0063] Step S502: Calculate the average value of the energy data collected by each group of dynamic sensors, and compare the calculated energy average value with the expected energy. In the case where the energy average value is in the first interval, replace the buffer cushion.

[0064] Specifically, under the action of the buffer cushion 6, the dynamic sensor 7 arranged below the buffer cushion 6 monitors the impact energy data in real - time. After calculating the average value of the data monitored by each group of dynamic sensors 7, the energy average value is obtained Q1 Then, compare Q 1 with the expected energy calculated in step S501. Q If the average energy Q 1 is in the first interval, i.e., 100% Q ≥ Q 1 ≥90% Q , it indicates that the buffer energy absorption material has failed and the buffer cushion needs to be replaced.

[0065] Step S503: When the average energy is in the second interval, compare the actual stress values collected by multiple groups of borehole stress gauges with the standard pressure, and determine the adjustment method of the buffer cushion according to the comparison result of the force data.

[0066] Specifically, if the average energy Q 1 is in the second interval, i.e., 90% Q ≥ Q 1 ≥70% Q , it indicates that the buffer material has an obvious buffering effect at this time. However, it is also necessary to consider the pressure changes collected in real time by the borehole stress gauges on the coal pillar side in the machine roadway and the air roadway. According to the average value F 1 of the actual stress values collected by 3 groups of borehole stress gauges, F compare it with the standard pressure

[0067] to determine the adjustment method. F ≥ F 1 ≥ F , keep the buffer cushion unchanged and no remedial measures are required; when the actual stress value is in the fifth interval, i.e., 1.5 F ≥ F 1 >1.2 F , it is necessary to increase the height of the buffer cushion, or other pressure relief measures can also be taken to reduce energy transfer; when the actual stress value is in the sixth interval, i.e., F 1 >1.5 F , it indicates that the buffer energy absorption material has failed and the buffer cushion needs to be replaced.

[0068] Among them, through the above examples, it can be seen that the sixth interval is greater than the fifth interval, and the fifth interval is greater than the fourth interval. The endpoints of each interval (i.e., the fourth interval, the fifth interval, and the sixth interval) are based on the standard pressure FDetermined.

[0069] Step S504, when the energy mean value is in the third interval, keep the buffer cushion layer unchanged.

[0070] Specifically, if the energy mean value Q 1 is in the third interval, that is, 70% Q ≥ Q 1 , it indicates that the performance of the buffer cushion layer 6 is good and it fully plays a buffering role, and no adjustment is required.

[0071] Among them, from the above description, it can be seen that the first interval is greater than the second interval, and the second interval is greater than the third interval. The endpoints of each interval (i.e., the first interval, the second interval, and the third interval) are determined according to the expected energy Q Determined.

[0072] Thus, this application prevents and controls the roof caving disaster by reducing the initial caving distance of the roof and slowing down the energy during roof caving, and can avoid accidents such as tipping of hydraulic supports and rock burst in the roadway due to roof caving.

[0073] In summary, for the method for preventing and controlling the roof caving disaster in the coal mining face of this application embodiment, in the side of the cut-eye of the coal mining face close to the coal pillar, pressure relief holes are drilled obliquely upward at a certain angle, and the hard rock layer of the roof is pre-cracked by hydraulic fracturing or blasting pressure relief, and the beam structure at the initial mining of the coal mining face is transformed into a cantilever structure. Thereby, the initial caving step distance of the roof can be effectively reduced, the over-large hanging roof area can be avoided, and the pressure manifestation degree of the working face can be reduced. Moreover, as the coal mining face advances, buffer materials are laid on the floor, and the excessive energy generated during roof caving is reduced to an acceptable range through the buffer materials, which can prevent various serious accidents caused by huge impact energy in the roadway and maintain the stability of the roadway. Thus, this method comprehensively prevents and controls the roof caving disaster from multiple angles, significantly reduces the impact caused by the roof caving disaster, and improves the reliability, pertinence, and accuracy of preventing and controlling the roof caving disaster in the coal mining face, which is beneficial to ensuring safe and stable mining in the mine.

[0074] Based on the above embodiments, in order to more clearly illustrate the specific implementation process of pre-cracking the hard rock layer of the roof by using the hydraulic fracturing process in this application by drilling fracturing holes, in a specific embodiment of this application below, in combination with Figure 3 and Figure 4 shown in the two hydraulic fracturing methods are used for exemplary illustration.

[0075] In this embodiment, it is required that the vertical height of the hydraulic roof cutting and pressure relief borehole be greater than the cable bolt support length of the cut-through roadway. For example, the vertical height of the borehole is generally greater than 6 m. Thus, even after pressure relief, the support system still plays a supporting role to ensure that the support system of the cut-through roadway does not fail. Moreover, in order to ensure the connection between the prefabricated fractures 4, when drilling the boreholes, the boreholes are grooved according to the fracturing position, and a two-way grooving method is adopted to ensure that the fractures are along the center line between the boreholes and the hydraulic fracturing propagation direction is consistent with the direction between the boreholes.

[0076] As a first example, when the vertical height from the construction position to the hard roof of the cut-through roadway (generally with a uniaxial compressive strength greater than 20 MPa) is less than 20 m, a short-hole fracturing scheme is adopted. The parameters that need to be designed mainly include the distance, angle between the short-hole fracturing borehole 5 and the coal pillar strike, and the distance between the boreholes.

[0077] For example, the distance between the boreholes is set not to exceed 6 m. Generally, the borehole is 2 m away from the coal pillar, which is convenient for the drilling rig to drill the borehole. The borehole length is 30 m, the elevation angle is 50°, and the included angle with the coal pillar is 0 - 10°. When drilling, it is drilled towards the direction of the working face. When fracturing to 8 m from the hole mouth, stop fracturing. The backward fracturing method is adopted during fracturing, and fracturing is carried out every 3 m. Stop when the pressure of the water pressure gauge decreases. The scenario after the short-hole fracturing borehole is drilled is as Figure 6 shown.

[0078] As a second example, when the vertical height from the construction position to the hard roof of the cut-through roadway (generally with a uniaxial compressive strength greater than 20 MPa) is greater than 20 m, a long-hole fracturing scheme is adopted. When drilling the fracturing borehole, ensure that the borehole is always located in the hard rock formation. There are two optional schemes in this example. The first scheme is as Figure 4 shown. Two horizontal long-hole fracturing boreholes 8 are drilled in the roof. Among them, the position of one borehole is at a distance of a1 (such as 1 m) from the coal pillar in the horizontal direction and at a distance of b1 (such as 1 m) from the coal pillar in the vertical direction. The position of the other borehole is at a distance of a2 (such as 3 m) from the coal pillar in the horizontal direction and at a distance of b2 (such as 20 m) from the coal pillar in the vertical direction. The second scheme is based on the first scheme. Along the side of the first coal pillar 1 in the machine roadway or the air roadway, drill a borehole parallel to the length of the cut-through roadway.

[0079] In this example, the backward fracturing method is also adopted, and stop when the pressure decreases. Fracturing is carried out every 5 m of backward movement. The two boreholes start from the side of the first coal pillar 1 in the cut-through roadway and are drilled upward in a stepped shape towards the working face side. Of course, other numbers of fracturing boreholes can also be drilled according to actual needs.

[0080] It should be noted that in the above first example and the second example, after the drilling construction is completed, it is necessary to slot towards the drilling positions on both sides at the hydraulic fracturing position respectively, and two-way slotting is required between the drill holes. For example, as Figure 7 shown, the directions of the slot openings 51 in the two short-hole fracturing drill holes 5 are opposite. This can ensure that the fracturing cut seam extends towards the cut-eye direction, and the cracks between the drill holes are completely penetrated.

[0081] To implement the above embodiments, the present application also proposes a prevention and control system for the roof caving disaster in the coal mining face. Figure 8 As shown in the structural schematic diagram of a prevention and control system for the roof caving disaster in the coal mining face proposed in the embodiment of the present application, as Figure 8 shown, the system includes: an acquisition module 100, a pre-fracture module 200, a setting module 300, and an adjustment module 400.

[0082] Among them, the acquisition module 100 is used to obtain the geological information of each rock stratum corresponding to the cut-eye of the coal mining face by drilling peepholes.

[0083] The pre-fracture module 200 is used to, based on the geological information, drill a plurality of pressure relief drill holes on the side of the first coal pillar corresponding to the cut-eye, and pre-fracture the hard roof rock stratum through pressure relief treatment to convert the roof at the cut-eye from a beam structure to a cantilever structure.

[0084] The setting module 300 is used to set a variety of monitoring devices at different positions in the air roadway, the machine roadway, and the goaf, and set multiple groups of buffer cushions composed of buffer energy-absorbing materials on the floor of the goaf. Among them, the height of the buffer cushion is determined based on the geological information, and the buffer cushion is used to relieve the dynamic pressure generated when the roof collapses to the floor.

[0085] The adjustment module 400 is used to combine the impact data collected by a variety of monitoring devices during roof caving, the initial caving step distance of the roof under the cantilever structure, and the relevant parameters of the cut-eye to determine whether to adjust the buffer cushion.

[0086] Optionally, in an embodiment of the present application, the setting module 300 is specifically used to: respectively set multiple groups of borehole stress gauges at the intermediate positions between the second coal pillar corresponding to the machine roadway and the third coal pillar corresponding to the air roadway, and set the standard pressure of the borehole stress gauges; set multiple groups of dynamic sensors along the parallel direction of the cut-eye, where each group of dynamic sensors includes a plurality of dynamic sensors arranged along the vertical direction of the cut-eye, and the dynamic sensors are used to monitor the impact energy generated during roof caving.

[0087] Optionally, in an embodiment of the present application, the adjustment module 400 is specifically configured to: calculate the expected energy generated when the roof collapses according to the initial caving step distance of the roof plate under the cantilever structure, the relevant parameters of the cut-through roadway, and the geological information; calculate the average value of the energy data collected by each group of the dynamic sensors, and compare the calculated energy average value with the expected energy. In the case where the energy average value is in the first interval, replace the buffer cushion layer; in the case where the energy average value is in the second interval, compare the actual stress values collected by the multiple groups of borehole stress gauges with the standard pressure, and determine the adjustment method of the buffer cushion layer according to the comparison result of the force data; in the case where the energy average value is in the third interval, keep the buffer cushion layer unchanged; wherein, the first interval is greater than the second interval, the second interval is greater than the third interval, and the endpoints of each interval are determined according to the expected energy.

[0088] Optionally, in an embodiment of the present application, the setting module 300 is specifically configured to: calculate the height of the initial roof caving based on the geological information; determine the height of the buffer cushion layer according to the flow parameters of the buffer energy-absorbing material and the height of the initial roof caving.

[0089] Optionally, in an embodiment of the present application, the pre-splitting module 200 is specifically configured to: drill a plurality of short-hole fracturing boreholes on the side of the first coal pillar corresponding to the cut-through roadway according to the strata strike information of the immediate roof and the main roof above the cut-through roadway; drill a plurality of short-hole fracturing boreholes on the side of the second coal pillar and the third coal pillar according to the geological information of the roof above the second coal pillar and the third coal pillar.

[0090] Optionally, in an embodiment of the present application, the pre-splitting module 200 is further configured to: set a single row of hydraulic props on the side of the first coal pillar corresponding to the cut-through roadway to support the roof at the cut-through roadway, wherein the distance between two adjacent hydraulic props in the single row of hydraulic props is the same as the bolt support spacing.

[0091] Optionally, in an embodiment of the present application, the acquisition module 100 is specifically configured to: drill a peephole perpendicular to the roof at a preset distance along the driving direction of the cut-through roadway, wherein the depth of the peephole is greater than twice the length of the roof cable bolt; draw a strata histogram of the cut-through roadway according to the peephole information obtained from each peephole.

[0092] It should be noted that the above explanation of the embodiments of the prevention and control method for the roof caving disaster in the coal mining face also applies to the system of this embodiment, and will not be elaborated here.

[0093] In summary, the prevention and control system for the roof caving disaster in the coal mining face according to the embodiments of the present application can effectively reduce the initial caving step distance of the roof, avoid excessive hanging roof area, and reduce the pressure manifestation degree of the working face. Moreover, the excessive energy generated during roof caving is reduced to an acceptable range through the buffer material, which can prevent various serious accidents caused by huge impact energy in the roadway and maintain the stability of the roadway. Therefore, the system comprehensively prevents and controls the roof caving disaster from multiple perspectives, significantly reducing the impact caused by the roof caving disaster.

[0094] To implement the above embodiments, the present application also proposes a non-transitory computer-readable storage medium, on which a computer program is stored. When the computer program is executed by a processor, it implements the prevention and control method for the roof caving disaster in the coal mining face as described in any one of the embodiments in the first aspect above.

[0095] In the description of this specification, the descriptions with reference to the terms "one embodiment", "some embodiments", "example", "specific example", or "some examples", etc. mean that the specific features, structures, materials, or characteristics described in connection with the embodiment or example are included in at least one embodiment or example of the present application. In this specification, the schematic representations of the above terms do not necessarily refer to the same embodiment or example. Moreover, the specific features, structures, materials, or characteristics described can be combined in a suitable manner in any one or more embodiments or examples. In addition, without contradiction, those skilled in the art can combine and combine the different embodiments or examples described in this specification and the features of different embodiments or examples.

[0096] In addition, the terms "first" and "second" are only used for descriptive purposes and cannot be understood as indicating or implying relative importance or implicitly specifying the quantity of the indicated technical features. Thus, the features defined with "first" and "second" can explicitly or implicitly include at least one of the features. In the description of the present application, the meaning of "a plurality" is at least two, such as two, three, etc., unless otherwise specifically defined.

[0097] Any process or method description shown in the flowchart or described in other ways herein can be understood as representing a module, segment, or part of the code including one or more executable instructions for implementing a customized logic function or process. The scope of the preferred embodiments of the present application includes additional implementations, where the functions can be executed in a substantially simultaneous manner or in an opposite order according to the involved functions, rather than in the order shown or discussed, which should be understood by those skilled in the art to which the embodiments of the present application belong.

[0098] The logic and / or steps represented in the flowchart or otherwise described herein can, for example, be considered a definitional sequence list of executable instructions for implementing logical functions, which can be embodied specifically in any computer-readable medium for use by an instruction execution system, apparatus, or device (such as a computer-based system, a system including a processor, or other systems that can fetch and execute instructions from the instruction execution system, apparatus, or device), or used in conjunction with these instruction execution systems, apparatuses, or devices. For the purposes of this specification, a "computer-readable medium" can be any device that can contain, store, communicate, propagate, or transport a program for use by or in conjunction with an instruction execution system, apparatus, or device. More specific examples (non-exhaustive list) of computer-readable media include the following: an electrical connection portion having one or more wirings (electronic device), a portable computer diskette (magnetic device), a random access memory (RAM), a read-only memory (ROM), an erasable programmable read-only memory (EPROM or flash memory), an optical fiber device, and a portable compact disc read-only memory (CDROM). Additionally, the computer-readable medium can even be paper or other suitable media on which the program can be printed, as the program can be obtained electronically, for example, by optically scanning the paper or other media, followed by editing, interpretation, or other suitable processing as necessary, and then stored in a computer memory.

[0099] It should be understood that various parts of the present application can be implemented by hardware, software, firmware, or a combination thereof. In the above embodiments, multiple steps or methods can be implemented by software or firmware stored in a memory and executed by a suitable instruction execution system. For example, if implemented in hardware, as in another embodiment, any one or a combination of the following techniques well known in the art can be used: discrete logic circuits having logic gate circuits for implementing logical functions on data signals, application-specific integrated circuits having suitable combinational logic gate circuits, programmable gate arrays (PGAs), field-programmable gate arrays (FPGAs), etc.

[0100] Those of ordinary skill in the art of this technology can understand that all or part of the steps carried by the method of implementing the above embodiments can be completed by a program instructing relevant hardware, and the program can be stored in a computer-readable storage medium. When the program is executed, it includes one or a combination of the steps of the method embodiments.

[0101] In addition, each functional unit in various embodiments of the present application may be integrated into a processing module, may exist physically alone for each unit, or two or more units may be integrated into one module. The above-mentioned integrated module may be implemented in the form of hardware or in the form of a software functional module. When the integrated module is implemented in the form of a software functional module and sold or used as an independent product, it may also be stored in a computer-readable storage medium.

[0102] The above-mentioned storage medium may be a read-only memory, a magnetic disk, an optical disc, etc. Although the embodiments of the present application have been shown and described above, it can be understood that the above embodiments are exemplary and should not be construed as limiting the present application. Those of ordinary skill in the art can make changes, modifications, substitutions, and variations to the above embodiments within the scope of the present application.

Claims

1. A method for preventing and controlling roof collapse disasters in a coal mining working face, characterized in that: The following steps are involved: By drilling peepholes, the geological information of each rock layer corresponding to the cut of the mining face can be obtained; Based on the geological information, a plurality of pressure relief boreholes are drilled on the side of the first coal pillar corresponding to the cutting eye, and the hard rock layer of the roof is pre-cracked by pressure relief treatment, so as to convert the roof at the cutting eye from a beam-like structure to a cantilever structure; A variety of monitoring equipment is set up at different locations of the air tunnel, machine tunnel and goaf, and a plurality of groups of buffer cushions composed of buffer energy absorbing materials are set up on the bottom plate of the goaf, wherein the height of the buffer cushion is determined based on the geological information, and the buffer cushion is used to alleviate the dynamic pressure generated when the roof collapses to the bottom plate; Based on the impact data of roof collapse collected by the various monitoring devices, the initial collapse step of the roof under the cantilever structure and the relevant parameters of the cutting eye, it is determined whether to adjust the buffer cushion layer.

2. The method according to claim 1, characterized in that A variety of monitoring equipment is set up at different locations of the ventilation tunnel, machine tunnel and goaf, including: A plurality of sets of borehole stress gauges are respectively arranged at the middle positions of the second coal pillar corresponding to the machine lane and the third coal pillar corresponding to the ventilation lane, and the standard pressures of the borehole stress gauges are set; A plurality of groups of dynamic sensors are arranged along the parallel direction of the cut-eye, wherein each group of dynamic sensors includes a plurality of dynamic sensors arranged along the vertical direction of the cut-eye, and the dynamic sensors are used to monitor the impact energy generated when the roof collapses.

3. The method according to claim 2, characterized in that The method of combining the impact data of the roof collapse collected by the multiple monitoring devices, the initial collapse step of the roof under the cantilever structure and the relevant parameters of the cut-eye to determine whether to adjust the buffer cushion layer includes: Calculating the expected energy generated when the roof collapses according to the initial collapse step of the roof under the cantilever structure, the relevant parameters of the cut-eye and the geological information; Calculating an average value of energy data collected by each group of the dynamic sensors, and comparing the calculated energy average value with the expected energy, and replacing the buffer pad layer when the energy average value is within a first interval; When the energy mean value is in the second interval, comparing the actual stress values ​​collected by the multiple groups of borehole stress gauges with the standard pressure, and determining the adjustment method of the buffer layer according to the force data comparison result; When the energy mean is in a third interval, the buffer layer is maintained unchanged; wherein the first interval is larger than the second interval, the second interval is larger than the third interval, and the endpoints of each interval are determined according to the expected energy.

4. The method according to claim 3, characterized in that The step of determining the adjustment method of the buffer layer according to the force data comparison result includes: When the actual stress value is within the fourth interval, maintaining the buffer layer unchanged; When the actual stress value is in the fifth interval, increasing the height of the buffer layer; When the actual stress value is in the sixth interval, the buffer pad layer is replaced; wherein the sixth interval is larger than the fifth interval, the fifth interval is larger than the fourth interval, and the endpoint of each interval is determined according to the standard pressure.

5. The method according to claim 1, characterized in that The method comprises setting a plurality of groups of buffer cushion layers composed of buffer energy absorbing materials on the bottom plate of the goaf, including: Calculate the height of the initial collapse of the roof based on the geological information; The height of the buffer cushion layer is determined according to the flow parameters of the buffer energy absorbing material and the height of the initial collapse of the top plate.

6. The method according to claim 2, characterized in that The pressure relief treatment includes hydraulic fracturing pressure relief and blasting pressure relief, and the pressure relief treatment is used to pre-crack the hard rock formation on the roof, including: According to the rock formation trend information of the immediate roof and the old roof above the cutting eye, a plurality of short-hole fracturing drill holes are drilled on the side of the first coal pillar corresponding to the cutting eye; According to the geological information of the roof above the second coal pillar and the third coal pillar, a plurality of short-hole fracturing drill holes are drilled near the second coal pillar side and the third coal pillar side.

7. The method according to claim 1, characterized in that After the pre-cracking of the hard rock formation on the roof by the pressure relief treatment, the method further comprises: A single row of hydraulic props is arranged on the side of the first coal pillar corresponding to the cutting eye to support the roof at the cutting eye, wherein the spacing between two adjacent hydraulic props in the single row of hydraulic props is the same as the spacing of anchor rod supports.

8. The method according to claim 1, characterized in that The peephole is drilled to obtain geological information of each rock layer corresponding to the cut of the mining face, including: According to the excavation direction of the eye cutting, a peephole perpendicular to the roof is drilled at a preset distance, wherein the depth of the peephole is greater than 2 times the length of the roof anchor cable; A rock formation column diagram of the cut eye is drawn according to the peep information obtained by each peep hole.

9. A system for preventing and controlling roof collapse disasters at a coal mining face, characterized in that: include: An acquisition module is used to obtain geological information of each rock layer corresponding to the cut of the mining face by drilling a peephole; A pre-splitting module is used to drill a plurality of pressure relief boreholes on the side of the first coal pillar corresponding to the cutting eye based on the geological information, and pre-splitting the hard rock layer of the roof through pressure relief treatment, so as to convert the roof at the cutting eye from a beam-like structure to a cantilever structure; A setting module is used to set a variety of monitoring equipment at different locations of the air tunnel, the machine tunnel and the goaf, and to set a plurality of buffer cushions composed of buffer energy absorbing materials on the bottom plate of the goaf, wherein the height of the buffer cushion is determined based on the geological information, and the buffer cushion is used to alleviate the dynamic pressure generated when the roof collapses to the bottom plate; The adjustment module is used to determine whether to adjust the buffer layer based on the impact data of the roof collapse collected by the various monitoring devices, the initial collapse step of the roof under the cantilever structure and the relevant parameters of the cutting eye.

10. A non-transitory computer-readable storage medium having a computer program stored thereon, characterized in that: When the computer program is executed by a processor, a method for preventing and controlling roof collapse disasters in a coal mining face as described in any one of claims 1 to 8 is implemented.

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