Method for preventing and controlling mine earthquake and rock burst hybrid power disaster of extremely thick hard rock downhill coal pillar

Abstract

The invention provides a mine earthquake and rock burst hybrid power disaster prevention and control method for an extremely thick hard rock downhill coal pillar, and relates to the technical field of mine earthquake disaster prevention and control. According to the method, a systematic prevention and control system is established by organically combining a plurality of links such as ground surface settlement observation and reasonable propulsion length determination, impact danger degree evaluation and region division, hybrid power impact monitoring and hybrid power prevention and control danger solving measures and comprehensively considering the coupling effect of a plurality of factors such as geological structures, mining technologies and stress distribution. By optimizing the reasonable advancing length of the working face, accurately dividing dangerous areas and implementing all-directional monitoring and targeted pressure relief measures, the occurrence risk of mine earthquakes and rock burst disasters is effectively reduced, and the safety of coal mining is guaranteed. The method provided by the invention achieves a remarkable prevention and control effect in practical application, and has important popularization and application values.

Description

Technical Field

[0001] The present invention relates to the technical field of mine earthquake disaster prevention and control, and in particular to a method for preventing and controlling composite dynamic disasters of mine earthquake and rock burst caused by huge thick hard rock downhill coal pillars. Background Art

[0002] Currently, mine-induced earthquakes are primarily prevented by weakening the ground structure to transform the thick, hard roof. Key methods include deep-hole blasting (DHB) and hydraulic fracturing. Li Junping et al. have further proposed controlled blasting and caving technology for thick ore bodies, which can release and divert high ground pressure.

[0003] In terms of theoretical research, Lu et al. studied the damage range of single-hole blasting of concentrated cylindrical charges based on a particle flow numerical simulation model to clarify the mechanism of DHB in controlling hard rock mine tremors. Dang et al. constructed a mechanical model of the interaction between the cantilever beam and support of the hard roof in the mining area, and obtained the support bearing capacity of different positions of the roof during the cyclic rotation of the working face, so as to achieve full support and safe and efficient mining of the working face. Li et al. established a mechanism model of hydraulic support resistance and direct coverage, and obtained that pre-cracking and cutting the roof can effectively control the mine pressure of thick hard roof. Zheng et al. studied and analyzed the fracture characteristics of the collapsed covering layer of the fully mechanized mining roof with hard roof and high mining height, and proposed a hard roof staged fracturing and advanced weakening technology.

[0004] In terms of rock burst prevention and control, Qi Qingxin et al. proposed the idea of preventing and controlling rock burst at the source in coal mines based on the basic theory of rock burst. Zhu Sitao et al., based on different types of mine tremors, revealed the mechanism by which isolated coal pillars of different widths induce local shock, overall shock, and mine tremors, and proposed a reasonable design method for the width of isolated coal pillars based on the coordinated control of rock burst and mine tremors. Zhai Minghua et al. studied the characteristics of rock burst under extremely thick and hard rock formations, revealing that it has three common characteristics: "critical working face effect", "vibration-induced shock effect", and "impact vibration effect". Based on the superposition principle of elastic mechanics, combined with the calculation of internal stress of components in rock mechanics and material mechanics, Yang Zhuang et al. derived the three-dimensional stress conversion formula at any position inside the coal pillar, providing a basis for the calculation of the maximum and minimum intermediate principal stresses.

[0005] In engineering practice, Sun Liuwei et al. used seismic wave CT technology to assess the risk of rock bursts in wide coal pillars and proposed a static load diversion anti-rock burst technology, including measures such as large-diameter drilling to relieve pressure, limiting daily advance, and strengthening support across the entire tunnel section. Yang Weili et al. proposed criteria for determining the overall instability of isolated working faces in thick igneous rock. They recommended designing the isolated working face as a wide-width island face for mining, thereby shifting the overall instability impact to localized impacts and reducing the overall instability risk. Li Hongyan et al., focusing on three aspects: coal rock properties, environmental stresses, and coal rock structure control, pointed out that efficient and intelligent prevention and control equipment is the only way to develop rock burst prevention and control technology. They also emphasized that integrated prevention and control of coal and rock dynamic hazards is an inevitable requirement for deep mining. Liu Shaohong et al. proposed a targeted coordinated prevention and control strategy based on the instability of the basic mechanisms of the complex mechanical system in the surrounding rock of large coal seam tunnels under the influence of mining and its sensitivity to surrounding environmental disturbances. Wang Peng et al., through engineering verification, concluded that staggered tunnel layout can address the problems of intense, frequent, and large deformation rock bursts in return air level tunnels. Based on the propagation characteristics of coal rock vibration energy, Wang Ning et al. proposed a genetic algorithm for large-scale station network layout combination planning problem - CMEAS algorithm.

[0006] In summary, research on the prevention and control of mine tremors and rock bursts has gradually moved from theoretical exploration to engineering practice, forming a comprehensive prevention and control system that spans source control, precise early warning, and regional prevention and control to localized pressure relief. However, with the increasing depth of coal mines, the prevention and control of mine tremors and rock bursts still faces numerous challenges. These include the randomness of mine tremor occurrence times and the uncertainty of the impact zones induced by mine tremors, which makes it difficult to accurately identify prevention and control locations. Further integration of theoretical research and engineering applications is needed to achieve more efficient and precise prevention and control goals. Summary of the Invention

[0007] In view of this, the present invention proposes a method for preventing and controlling composite dynamic disasters of mine earthquake and rock burst in thick hard rock downhill coal pillars to solve the problems existing in the above-mentioned prior art.

[0008] To achieve the above objectives, the present invention proposes a method for preventing and controlling the combined dynamic disaster of mine earthquake and rock burst in thick hard rock downhill coal pillars, comprising:

[0009] Set up observation lines on the coal pillar working face to obtain surface settlement curves;

[0010] Determine the reasonable advancement length of the coal pillar working face based on the surface settlement curve and the numerical simulation results;

[0011] The comprehensive index method is used to evaluate the impact hazard degree of the coal pillar working face and determine the rock burst hazard level;

[0012] Dividing regions based on the rock burst hazard level;

[0013] Conduct composite power impact monitoring in the divided areas and formulate composite power prevention and control measures.

[0014] Furthermore, the process of dividing the regions based on the rock burst hazard level includes:

[0015] The multi-factor coupling evaluation method is used to analyze the impact factors of rock burst in the divided area. Based on the original rock deadweight stress field, the stress field superposition principle is used to synthesize the additional stress generated by tectonic stress, mining stress or other disaster-causing factors to obtain the total stress field strength of the coal rock.

[0016] The rock burst hazard level areas are divided based on the ratio of the total stress field intensity to the uniaxial compressive strength.

[0017] Furthermore, the composite dynamic impact monitoring includes regional impact monitoring and local impact monitoring, and the local impact monitoring includes coal stress monitoring, roadway displacement monitoring, and support pressure monitoring;

[0018] The regional impact monitoring process includes: arranging several groups of microseismic monitoring points along the direction of mining advancement, with the distance between adjacent monitoring points controlled within 200m to 300m, issuing an early warning when the daily cumulative energy detected exceeds 3.5E+05J; issuing an early warning when the energy release of a single vibration event reaches 2.0E+05J; and issuing an early warning when the density of high-frequency vibration events exceeds 5 times within 24 hours.

[0019] Furthermore, the coal body stress monitoring includes: arranging stress gauges in a double-hole gradient, with each group of stress gauges including two measuring points, one at a shallow hole of 12m and one at a deep hole of 18m, with the spacing of the measuring points within the group along the roadway strike ≤2m, and the spacing of adjacent monitoring groups ≤30m, and continuously arranged along the roadway strike, with the first group no more than 40m away from the working face, and installed on the east side of the transport chute, the west side of the return air chute, the east side of the auxiliary return air chute, and the upper side of the auxiliary cutting hole, at a distance of 0.5m to 1.5m from the roadway floor, and moving forward with the working face mining;

[0020] The tunnel displacement monitoring includes installing a set of roof separation and tunnel deformation monitoring equipment every 100m in the transport chute and return air chute;

[0021] The support pressure monitoring includes installing a pressure monitoring station every three supports in the three-dimensional coal pillar area, and arranging them at a spacing of 10 supports in the extended coal pillar working face area. The periodic pressure situation of the working face is predicted based on the initial collapse step distance and periodic fracture of the direct roof and old roof. Before the working face is mined, the integrity index of the roof is counted and the activity pattern of the old roof is judged.

[0022] Furthermore, the composite power prevention and control measures include pre-pressure unloading measures during excavation and pressure unloading measures during mining. The pre-pressure unloading measures during excavation include pre-pressure unloading by large-diameter drilling on the side of the tunnel and pre-pressure unloading by breaking the bottom during excavation. The pressure unloading measures during mining include pre-pressure unloading by large-diameter drilling on the side of the tunnel, breaking the bottom measures and breaking the top measures.

[0023] Furthermore, the large-diameter drilling of pre-pressure relief holes on the roadway side during excavation includes: drilling pressure relief holes perpendicular to the coal wall on both sides of the return air chute and the transport chute roadway, not constructing pressure relief holes on the goaf side where the coal pillar is less than 15m, and constructing pressure relief holes on the goaf side where the coal pillar is greater than 15m;

[0024] Pressure relief holes are drilled on the upper side of the auxiliary cutting eye, but not on the lower side;

[0025] Drill pressure relief holes on the vertical coal wall of the west side of the auxiliary return air chute, drill pressure relief holes below 140m on the east side of the auxiliary return air chute, and no pressure relief holes will be drilled within the range of 0-140m;

[0026] Pressure relief holes are drilled on the upper side of the cut eye, but not on the lower side.

[0027] Furthermore, the pre-pressure relief of the broken bottom includes: taking floor pressure relief measures when the thickness of the bottom coal on one side of the roadway is greater than 1m;

[0028] For thin low coal with thickness d satisfying 1m<d≤1.5m, pressure relief holes / water injection holes are arranged at the low angles of the double sides, and water is injected alternately through every other hole;

[0029] For medium-thick low-lying coal with a thickness d of 1.5 < d ≤ 4m, water injection holes and blasting holes are constructed alternately at the low angles of the double sides, and high-pressure water injection is performed on the water injection holes;

[0030] For thick low-lying coal with thickness d>4m, two bottom blasting holes are added every 10-20m on the basis of the pressure relief measures for medium-thick low-lying coal;

[0031] No bottom blasting is carried out on the crosshead and the tunnel floor within 3m from the crosshead. Only bottom pressure relief holes and bottom water injection holes are constructed. The hole spacing is adjusted to 1m, and water is injected alternately through the holes.

[0032] Furthermore, the pressure relief measures for large-diameter drilling holes on the side of the roadway during mining include:

[0033] On the basis of pre-unloading of the roadway side with large diameter drilling holes during the excavation, pressure relief holes are drilled at intervals of 1.5m in the high stress area of the cutting eye, and double-stage drilling is implemented on the east side within the 300m advance range of the transport drift advance area. For the auxiliary cutting eye and return air drift, inclination gradient control and spacing graded control are used to unload the heterogeneous coal body. For the return air drift, penetrating holes are drilled on the west side and the pressure relief range of the advance working face is not less than 150m.

[0034] Furthermore, the bottom breaking measures include: on the basis of the pre-unloading of the bottom coal, strengthening the monitoring of the bottom coal during the mining period, and continuing to take bottom breaking measures if the bottom coal still has the risk of impact. The bottom breaking parameters are based on the bottom breaking parameters during excavation, and are adjusted in time when the anti-impact requirements are not met.

[0035] Furthermore, the roof breaking measures include: drilling holes at intervals of 8m along the side roof of the coal pillar on the empty coal pillar greater than 10m in the working face transport chute and return air chute, and positioning the final hole at the mudstone-sandstone interlayer interface.

[0036] Compared with the prior art, the present invention has the following beneficial effects:

[0037] This example systematically illustrates the process of dynamic disaster prevention and control based on the combined dynamic mechanism of mine tremors and rock bursts. By optimizing design and delineating control areas, combined with monitoring, early warning, and mitigation measures, the method ultimately verifies its effectiveness. Ultimately, a technical system for combined dynamic disaster prevention and control of mine tremors and rock bursts in the lower coal pillar was established, effectively reducing the occurrence of disasters and ensuring safe production in coal mines. BRIEF DESCRIPTION OF THE DRAWINGS

[0038] Various other advantages and benefits will become apparent to those skilled in the art upon reading the detailed description of the preferred embodiment below. The accompanying drawings are for illustration purposes only and are not to be considered as limiting the present invention. The same reference symbols are used throughout the drawings to represent the same components. In the drawings:

[0039] Figure 1 This is a layout diagram of the surface settlement observation line of the 21010 working face in an embodiment of the present invention;

[0040] Figure 2 Surface settlement curves along the strike and dip directions during mining at the 21010 working face in an embodiment of the present invention, wherein (a) is the strike surface settlement curve and (b) is the dip surface settlement curve;

[0041] Figure 3 The plane vertical stress cloud diagrams during the advancement of the working face of Example 21010 of the present invention, where (a) is when the working face advances 150 m, (b) is when the working face advances 250 m, (c) is when the working face advances 350 m, and (d) is when the working face advances 450 m.

[0042] Figure 4 This is a 21010 working face roadway layout diagram in an embodiment of the present invention;

[0043] Figure 5 This is a schematic plan view of the division of rock burst hazard zones in the 21010 mining face roadway in an embodiment of the present invention;

[0044] Figure 6 Schematic diagram of the arrangement of microseismic measurement points in an embodiment of the present invention;

[0045] Figure 7 Schematic diagram of drilling arrangement for working face mining stress monitoring in an embodiment of the present invention;

[0046] Figure 8 Schematic diagram of the bottom pressure relief of the 21010 working face in an embodiment of the present invention, wherein (a) is a plan view of the bottom pressure relief, and (b) is a schematic diagram of the II section;

[0047] Figure 9 Schematic diagram of deep hole blasting drilling in the roof according to an embodiment of the present invention, wherein (a) is a schematic diagram of the planar layout, and (b) is a schematic diagram of the cross-section along the working face;

[0048] Figure 10 This is the daily microseismic frequency and total energy change curve from March to July in the embodiment of the present invention;

[0049] Figure 11 The stress change curves of different measuring stations from March to April in the embodiment of the present invention, where (a) is measuring station 7, (b) is measuring station 31, (c) is measuring station 33, and (d) is measuring station 35;

[0050] Figure 12 The stress change curves of different anchor rods from March to April in the embodiment of the present invention are shown, where (a) is anchor rod No. 1, (b) is anchor rod No. 3, (c) is anchor rod No. 5, (d) is anchor rod No. 7, and (e) is anchor rod No. 9;

[0051] Figure 13 Schematic diagram of the overall process of the method of the present invention. DETAILED DESCRIPTION

[0052] The exemplary embodiments of the present disclosure will be described in more detail below with reference to the accompanying drawings. Although exemplary embodiments of the present disclosure are shown in the accompanying drawings, it should be understood that the present disclosure can be implemented in various forms and should not be limited by the embodiments set forth herein. On the contrary, these embodiments are provided to enable a more thorough understanding of the present disclosure and to fully convey the scope of the present disclosure to those skilled in the art. It should be noted that, unless there is a conflict, the embodiments of the present disclosure and the features described in the embodiments can be combined with each other. The present invention will be described in detail below with reference to the accompanying drawings and in combination with the embodiments.

[0053] This embodiment proposes a method for preventing and controlling combined dynamic disasters of mine earthquakes and rock bursts caused by extremely thick hard rock downhill coal pillars, including:

[0054] 21010 Determination of reasonable length of working face advancement:

[0055] Determining the working face advance length plays a crucial role in coal mining design and requires comprehensive consideration of multiple relevant influencing factors. This example mainly analyzes the following three aspects:

[0056] From the perspective of surface settlement around the working face, it can be seen from the surface settlement data. The west side of the coal pillar in the 21 mining area is a large-scale goaf area, and its settlement trend gradually increases from the southern downhill area to the north, and the settlement value increases from the initial 2.26m to 7.5m; while the cumulative settlement value on the east side is relatively stable, and it is better to maintain it at around 5.7m. Due to the existence of goafs on both sides, surface settlement will increase the load on the coal pillar, generate higher vertical stress, and form a stress concentration area. When the coal pillar is under high stress for a long time, the strength of the coal pillar will decrease, and the risk of instability will easily occur. In addition, long-term high stress may also induce creep of the coal pillar, causing deformation to increase over time. In order to further explore the settlement of the coal pillar surface during the advancement of the working face, an observation line was set up on the surface strike and dip of the first mining working face of coal pillar 21010. The schematic diagram of the layout of the surface settlement observation line is shown as follows. Figure 1 Through observation, the surface settlement curves of the dip and strike from the initial mining period to June 9, 2023, July 14, 2023, and August 14, 2023 were obtained, as shown in Figure 2. Figure 2 shown.

[0057] exist Figure 2 It can be seen that as time goes by and the working face continues to advance, the surface settlement displacement is also increasing. Comparing the settlement curves from the beginning of mining to June 9, 2023 and August 14, 2023: Figure 2 As shown in (a), the surface settlement displacement is more obvious at the cut hole, and a local depression appears. As the distance from the cut hole increases, the surface settlement rate becomes more stable and the settlement value becomes smaller, indicating that the settlement distribution along the strike direction is relatively uniform, that is, the length of the working face determines the settlement range. Figure 2 As shown in (b), the settlement values along the working face (measurement points 11-18) are all around 0.25 mm along the dip direction. This is inconsistent with the conventional pattern of maximum settlement near the center of the goaf and gradually decreasing toward the sides. This indicates that the overlying thick hard rock layer is in the stage of accumulation of settlement energy. As the working face advances, the accumulated elastic potential energy continues to increase. Therefore, it is necessary to control the advancement length of the working face and relieve pressure by breaking the roof to release the accumulated elastic potential energy.

[0058] Based on the critical fracture span of the thick hard rock strata overlying the coal pillar, combined with numerical simulation results, it is estimated to be 550 to 600 meters. When the mining length does not reach the critical span, the rock strata bend and deform, storing a large amount of elastic energy within them. When this elastic energy accumulates to a certain level (the critical span), the rock strata fracture and generate microseismic events with enormous energy, inducing mine tremors, affecting the safety of surface buildings and potentially triggering rock bursts, posing a threat to mine safety. Therefore, it is necessary to optimize mining plans to avoid the occurrence of strong mine tremors.

[0059] Since the eastern part of the 21010 working face is solid coal, it can support the overlying hard rock layer during the coal pillar mining process, preventing the roof from sinking or collapsing, and has a certain protective effect. Through numerical simulation, it can be obtained that the plane vertical stress cloud map during the advancement of the 21010 working face (see Figure 3 ).exist Figure 3 It can be found that due to the supporting effect of solid coal on the overlying strata, it can effectively disperse and bear the pressure from the overlying strata, thereby reducing the vertical stress on the working face during the advancement process. In addition, solid coal helps to control the deformation of the working face and the surrounding rock strata, can buffer and absorb part of the stress, and reduce stress concentration. However, as the advancement distance increases, the protective effect of solid coal weakens, and the leading coal body on the transport chute side is in a state of stress concentration. Especially when the working face advances to the goaf on both sides, the goaf of the working face and the goaf on both sides are interconnected, and a high stress area is formed in the leading area of the working face, which is vulnerable to the threat of impact ground pressure. Therefore, the longer the working face advances, the weaker the protective effect of the solid coal in the eastern part of the working face, which is not conducive to maintaining the stability and safety of the working face.

[0060] Taking all the above factors into consideration, the advancement length of the 21010 working face should not be too long: during the advancement of the working face, it is necessary to constantly monitor the surface settlement to determine whether the thick overburden has moved; the advancement length of the working face must be controlled within the ultimate breaking span of the thick hard rock layer to avoid the sudden breaking of the hard rock layer causing high-energy strong mine earthquakes; at the same time, the advancement length of the working face must be within the protection range of the eastern solid coal to avoid the working face being in a high stress state for a long time and causing rock bursts. Combined with the actual situation of the on-site project, the final advancement length of the working face was determined to be 450m. Figure 4The total length of the return air chute and the auxiliary return air chute of the 21010 working face is 450m, the length of the transport chute is 450m, the length of the main cutting eye is 260m, the length of the auxiliary cutting eye is 60m, and the recoverable length is 450m.

[0061] 21010 working face during mining period, key prevention and control areas are delineated:

[0062] The comprehensive index method is currently the most commonly used method for evaluating rock burst risk. It mainly uses the geological factor evaluation index and the mining technology factor evaluation index to comprehensively evaluate the rock burst risk level in the mining area of the working face. t The calculation formula is as follows:

[0063] W t =max{W t1 ,W t2}(1)

[0064] Where: W t1 W is an index for evaluating the effect of geological structural factors on rock burst and the rock burst hazard level; t2 An index for evaluating the effect of mining technology factors on rock burst and the hazard level of rock burst.

[0065] Where W t1 , W t2 The calculation formula is:

[0066]

[0067] Where: W i is the actual index of the i-th influencing factor around the mining face; W i max is the maximum impact risk index of the i-th influencing factor; n is the number of influencing factors.

[0068] Using the comprehensive index method based on engineering data from the 21010 working face, the rock burst hazard assessment indices for geological and structural factors and mining technology were calculated to be 0.52 and 0.73, respectively. Combined with the four levels of rock burst hazard determined by the comprehensive index method (see Table 1), the rock burst hazard for the 21010 working face is rated C (moderate rock burst).

[0069] Table 1

[0070]

[0071] On the basis of the comprehensive index method, a multi-factor coupling evaluation method is adopted, that is, by analyzing the factors affecting rock burst in a certain area, taking the original rock deadweight stress field as the benchmark, the principle of stress field superposition is adopted to synthesize the additional stresses generated by disaster-causing factors such as tectonic stress and mining stress, and obtain the total stress field strength of coal rock. On this basis, an impact hazard criterion index is established (that is, the ratio of the total stress field strength of coal rock to its uniaxial compressive strength) to achieve a multi-level early warning division of rock burst hazard. According to the calculation results of engineering data, it is concluded that there are 2 weak impact (blue areas) and 5 medium impact (yellow areas) dangerous areas. The specific results are shown in Table 2. Figure 5 shown.

[0072] Table 2

[0073]

[0074]

[0075] 21010 working face composite dynamic impact monitoring:

[0076] Regional impact monitoring program:

[0077] The SOS microseismic monitoring system is used to monitor the complex dynamic disaster area of the 21010 mining face. The implementation plan is to arrange 6 groups of microseismic monitoring points along the mining advance direction, with the distance between adjacent monitoring points controlled within 200m to 300m, so as to achieve spatial coverage of the advanced support pressure zone. The layout diagram is shown in the figure below. Figure 5-6 As shown. When the distance between the shearer's cutting trajectory and the nearest monitoring point reaches a critical value (<100m), the monitoring point needs to be relocated promptly. During the initial mining period of the 21010 working face, the microseismic early warning indicators are temporarily determined based on the similar conditions of the final mining period of the 21170 working face. An early warning is issued when the monitoring system detects that the daily cumulative energy exceeds 3.5E+05J; an early warning is issued when the energy release of a single vibration event reaches 2.0E+05J; and an early warning is issued when the density of high-frequency vibration events (≥4th power) exceeds 5 within 24 hours.

[0078] Local impact monitoring solution:

[0079] (1) Coal stress monitoring

[0080] The MZ21010 mining face coal body stress monitoring uses the stress meter KJ743 installed during excavation to conduct real-time monitoring of the stress changes of the working face. The stress monitoring system is arranged according to a double-hole gradient. Each monitoring group contains two deep and shallow measuring points (shallow hole 12m, deep hole 18m). The spacing between the measuring points in the group along the direction of the tunnel is ≤2m, and the interval between adjacent monitoring groups is ≤30m and they are continuously arranged along the direction of the tunnel. The first group is no more than 40m away from the working face and is installed on the east side of the working face transport chute, the west side of the return air chute, the east side of the auxiliary return air chute, and the upper side of the auxiliary cutting eye. It is 0.5m to 1.5m away from the bottom plate of the tunnel. The group is withdrawn and moved forward in time as the working face is mined. The working face stress monitoring boreholes (green is deep hole, blue is shallow hole) are arranged as follows Figure 7 As shown, the installation spacing, hole depth, and number of monitoring points within a group are adjusted based on actual conditions. Single stress measuring points are deployed in the section with a coal pillar width of 10 to 17.6 meters, with an installation depth of half the coal pillar width. The initial critical stress warning indicator value is based on the depth of the measuring point: a warning is issued if the pressure is greater than or equal to 10 MPa in shallow holes (12 meters), and if it is greater than or equal to 12 MPa in deep holes (18 meters).

[0081] (2) Tunnel displacement monitoring

[0082] Roof separation and roadway deformation monitoring equipment will be installed in the transport and return air chute. A set of these monitoring devices will be installed every 100 meters on the 21010 working face, covering both the transport and return air chute. Since the roof of the auxiliary return air chute on the 21010 working face is located within the 21 extended goaf, roof separation sensors cannot be installed. Instead, only roadway deformation monitoring devices will be installed, with one set installed every 100 meters, covering the auxiliary return air chute.

[0083] (3) Bracket pressure monitoring

[0084] A KJ653 roof dynamic monitoring system is installed on the working face supports. A pressure monitoring station is installed every three supports in the 110.6m three-dimensional coal pillar area, and a pressure monitoring station is installed at intervals of 10 supports in the 21 extended coal pillar working face area. By judging the changes in real-time data collected by these monitoring stations, the initial collapse step distance of the direct roof and the old roof, as well as the periodic fracture situation, the cyclic pressure of the working face is predicted. Before the working face is mined, the integrity index of the roof (slab depth, caving volume) is calculated to assist in determining the activity pattern of the old roof, and the mining speed is strictly controlled according to microseismic and other monitoring data.

[0085] Composite power prevention and control measures for working face:

[0086] Currently, measures to prevent and control isolated coal pillars in thick, hard rock formations are divided into two aspects: prevention and mitigation. Preventing impacts involves implementing appropriate measures to prevent impact hazards before mining. Mitigation measures primarily include large-diameter drilling and water injection to relieve pressure on the coal body, and roof fracturing, blasting, and directional cracking to weaken the integrity of the hard roof.

[0087] Pre-pressure relief measures during excavation:

[0088] (1) Pre-unloading of large diameter drilling holes in the tunnel side

[0089] Pressure relief holes with a diameter of 125mm are drilled perpendicular to the coal wall on both sides of the return air chute and transport chute. The hole spacing is no more than 2m, and the hole depth is 25m. The drilling angles on both sides of the transport and return air chute are 5° to 12°. Since one side of the transport chute and return air chute is close to the goaf, a coal pillar of 0 to 17.6m is left. No pressure relief holes are drilled on the goaf side for coal pillars less than 15m. Pressure relief holes are drilled on the goaf side for coal pillars greater than 15m, with a hole depth of 10m.

[0090] The drilling angle of the upper side of the auxiliary cutting eye is 8°~12°. Since the 6m lower side of the auxiliary cutting eye is the 21170 lower side, no pressure relief holes will be constructed on the lower side (only the east side of the 21010 transport drift is used for pressure relief drilling, and only the west side of the 21010 return air drift 21170 lower side is used for pressure relief drilling upwards. The upper side of the solid coal body section of the auxiliary cutting eye is constructed).

[0091] Pressure relief holes with a diameter of 125mm are drilled in the vertical coal wall of the west side of the auxiliary return air chute, with a hole spacing of no more than 3m and a hole depth of 15 to 18m (the west side is 21 to 23m away from the 21st extension auxiliary tunnel, with a safety distance of 5m reserved). Pressure relief holes are constructed below 140m on the east side of the auxiliary return air chute, with a hole spacing of more than 3m and a hole depth of 25m (no pressure relief holes are constructed within the range of 0 to 140m on the east side).

[0092] Cut the eye (no pressure relief drilling will be constructed on the wall of the lower tunnel in the goaf). The drilling angle of the upper wall is 8°~12°, the hole spacing is not more than 2m, and it is 0.5~1.5m away from the bottom plate (the lower wall is only 10m away from the lower tunnel of 21220, and no pressure relief drilling will be constructed).

[0093] Detailed parameters of large-diameter pressure relief drilling holes during excavation are shown in Table 4.

[0094] Table 4

[0095]

[0096] (2) Pre-pressure relief measures for bottom break

[0097] Basic pressure relief principle: When the bottom coal thickness (d) on one side of a roadway exceeds 1m, floor pressure relief measures must be implemented. Based on field measurement data, the bottom coal thickness is divided into three levels, and targeted pressure relief solutions are implemented for each level.

[0098] ① Pressure relief measures for thin and low coal (1m<d≤1.5m): arrange 75mm diameter pressure relief holes / water injection holes at the low angles of both sides, with an inclination of 50°±10°, a hole spacing of 1m±0.2m, and alternate water injection through every other hole.

[0099] ① Medium-thickness low-coal pressure relief measures (1.5<d≤4m): alternately construct 75mm diameter water injection holes and 42mm diameter blasting holes at low angles on both sides, with an inclination of 50°±10° and a hole spacing of 1m±0.2m. Use the "one blast every other" process and high-pressure water injection in non-blasting holes.

[0100] ① Enhanced pressure relief measures for thick low-lying coal (d>4m). On the basis of medium-thickness measures, deep pressure relief of the bottom plate is added. Two bottom plate blasting holes are added in the cross section every 10-20m. The blasting holes are 42mm in diameter, with an inclination of 45°~90°, a spacing of 1-1.5m, and the first hole is 1m away from the wall.

[0101] ① Control measures for special areas: To ensure the strength and stability of the crosshead support, no bottom blasting is performed on the crosshead and the roadway floor within 3m of the crosshead. Only bottom pressure relief holes and bottom water injection holes are constructed. The hole spacing is adjusted to 1m, and water injection is carried out alternately. When passing through the old roadway, the roof is broken. To ensure the stability of the support, no bottom blasting is performed on the old roadway section and the roadway floor within 3m before and after. Only bottom pressure relief holes and bottom water injection holes are constructed. The hole spacing is adjusted to 1m, and holes are alternately constructed. See the schematic diagram of the bottom pressure relief of the working face for details. Figure 9 .

[0102] Pressure relief measures during mining:

[0103] (1) Pressure relief measures for large-diameter drilling holes in the tunnel side

[0104] This pressure relief is carried out on the basis of pre-pressure relief during tunneling. The diameter of the drill holes is uniformly 125mm, and the hole height is 0.5 to 1.5m from the bottom plate, with an error control within the range of ±0.2m. The drilling is carried out according to the principle of drilling 50m and then mining 30m. During the mining 30m, another round of pressure relief drilling is carried out. The detailed plan is as follows:

[0105] ① High stress area of the cut: along the 114m coal pillar on the west side, pressure relief drill holes with an inclination of 10 to 13°, a hole depth of 50m and a spacing of 1.5m are arranged to form a 30m dynamic pressure relief control area.

[0106] ① In the advance area of the haulage drift, a two-stage drilling pattern was implemented on the east side within the 300m advance range. During the foundation phase, 25m deep holes were arranged with an inclination of 5-12° and a spacing of 2m between holes. During the intensified mining phase, a round of pressure relief drilling was carried out from 50m upwards, with a depth of 50m and a spacing of 1m between holes.

[0107] ① Auxiliary cuthole and auxiliary return air chute: For both the auxiliary cuthole and the return air chute, gradient control technology is used, with an upper 8° to 12° angle and a side 5° to 12° angle. Drilling depths are designed to be 15 to 18 meters for the west chute and 20 and 25 meters for the east chute. Staggered spacing (1 meter for the auxiliary cuthole and 2 meters for the auxiliary return air chute) is used to achieve pressure relief and efficiency improvement in heterogeneous coal bodies.

[0108] ① Return air chute: On the west side, penetrating holes with an inclination of 5 to 7 degrees are used. The hole depth is 43m, the spacing is 1m, and the pressure relief range of the leading working face is not less than 150m.

[0109] The parameters of large-diameter pressure relief drilling during mining are shown in Table 5.

[0110] Table 5

[0111]

[0112]

[0113] The relevant requirements are as follows:

[0114] ①The 21010 mining face exposes many old tunnels. In order to prevent spontaneous combustion disasters caused by the communication between drilling holes and old tunnels, according to the requirements of the mine ventilation management system, the 21220 old lower tunnel, 21220 upper tunnel, and 21200 lower tunnel are orthogonal to the transport chute, and the 21170 lower tunnel is orthogonal to the auxiliary return air chute. No pressure relief drilling will be constructed in the old tunnel and the range of 6m before and after.

[0115] ① The depth of the transport chute, auxiliary return air chute, auxiliary cutting eye and return air chute shall not exceed 25m. The distance between the pressure relief drilling hole and the leading working face shall not be less than 300m.

[0116] ① Due to the presence of significant water in the abandoned roadway on the east side of the auxiliary return air chute within a 0-120m radius, the mine's geophysical survey department designated the abandoned roadway as a waterlogged area, prohibiting drilling within the water exploration line. The east side pressure relief borehole was designed to be 20m long to avoid the possibility of a water inrush channel created by the existing anti-scouring work on the west side of the original return air chute.

[0117] (2) Bottom-cutting measures

[0118] During the 21010 working face, floor-breaking measures are implemented based on pre-decompression of the bottom coal during tunneling. During the working face, floor-breaking measures are closely monitored. If the risk of impact continues, floor-breaking measures must be continued. Floor-breaking parameters are based on those used during tunneling. If impact prevention requirements are not met, they must be adjusted promptly.

[0119] (3) Top-cutting measures

[0120] The implementation area is the empty coal pillar greater than 10 meters in the haulage and return air chute of the 21010 working face. The drilling parameters are: drilling at an elevation angle of 83° ± 2° along the roof of the coal pillar, with the final hole positioned at the interface of the mudstone-conglomerate interbeds, and spacing 8 meters apart. Detailed parameters are shown in Table 6.

[0121] Table 6

[0122]

[0123]

[0124] Analysis of prevention and control effects on working face:

[0125] Analysis of microseismic event distribution patterns:

[0126] Since the coal pillar is of unequal thickness, no coal pillars with a thickness greater than 10 mm did not occur at the transport chute in March. 4 However, as the working face advanced, there were more than 10 microseismic events at the transport chute in April and May. 4 J microseismic events, and the frequency of microseismic events gradually increased. 4 There were 2 microseismic events in J, which increased to 4 in May, with the largest one reaching an energy of 8.7×10 4 J, immediate pressure relief measures must be taken. In terms of height, the development height of microseismic events is not obvious, and they are mainly concentrated in the 50m thick mudstone layer. This is mainly because the thick hard rock layer above is not easy to break, resulting in the collapse of the overburden and the development height of the cracks reaching the limit, and it cannot continue to expand upward. At this time, the thick hard rock layer is in a state of energy accumulation. Since June, the concentration of microseismic events in the transport chute has been significantly reduced, and the energy is greater than 10 4 The number of microseismic events in J was reduced to 0 again, proving the good effect of the pressure relief measures during the period.

[0127] Combined with the daily microseismic frequency and total energy change curves from March to July during the mining process of the 21010 working face Figure 10 As can be seen, the daily microseismic frequency and total energy curves in March and April both showed an upward trend as the working face advanced, corresponding to the accumulation and release of energy. However, when the curves began to rise sharply, the risk of rock bursts was high. The total energy curve reached a maximum of 1.5×10⁴J on May 18th, but the microseismic frequency was at its minimum at this time, indicating that energy was accumulating and requiring immediate decompression measures. Subsequently, in late May, the daily microseismic frequency showed an upward trend, while the total energy showed a downward trend, corresponding to the slow release of energy during the decompression process. This indicates that the decompression effect in May was significant and reduced the risk of rock bursts. After entering June, the two curves regained stability and became parallel, indicating that energy release was becoming more stable.

[0128] Mining stress analysis:

[0129] Taking the mining area in front of the 21010 working face as the research object, the 21010 mining area was fully surrounded by monitoring. The monitoring data of 10 stations, including No. 1, No. 3, No. 5, No. 7, and No. 9 of the return air chute and the auxiliary return air chute, and No. 31, No. 33, No. 35, No. 37, and No. 39 of the transport chute, were selected for analysis. The distances of these stations from the cutting eye are 424m, 374m, 326m, 287m, 429m, 380m, 334m, 357m, and 307m, respectively. The stress data of the shallow hole and deep hole stress gauges in March and April were extracted and plotted into a stress curve change diagram ( Figure 11 ).

[0130] Results show that stress changes at the return air chute and auxiliary return air chute stations were relatively stable, remaining below the warning line. At station 7, shallow hole stress values have been increasing since early March as the working face advanced, approaching the shallow hole warning line. Pressure relief measures implemented on March 14th moved stress values away from the shallow hole warning line, but as the working face continued to advance, stress values continued to rise, necessitating phased pressure relief to keep them below the shallow hole warning line. Stress at the transport chute stations remained low, remaining below the warning line. Pressure relief measures implemented at station 31 on April 12th further reduced deep hole stress values, moving them away from the warning line. Deep hole stress values at station 33 showed an increasing trend. To reduce stress concentration, pressure relief was implemented twice, on March 10th and April 20th, respectively, effectively controlling stress growth and keeping both shallow and deep hole stress values below the warning line. At station 35, shallow hole stress values have been increasing since pressure relief on March 7th as the working face advanced. In early April, due to rapid footage speeds and incomplete pressure relief, stress levels dropped sharply before rising again. During this period, shallow-hole stress values remained higher than deep-hole stress values, reaching a peak of 7.6 MPa on April 24. To prevent this hazard, enhanced pressure relief was immediately implemented. Starting on April 27, shallow-hole stress values rapidly dropped 47% to 3 MPa. Subsequently, stress values returned to normal, showing no upward trend and remaining consistently lower than deep-hole stress values. Both shallow and deep-hole stress values remained stable below the warning line. These analysis demonstrates that online dynamic stress monitoring and staged pressure relief can effectively mitigate the risk of sudden stress increases, achieving significant pressure relief results.

[0131] Analysis of the stress of tunnel anchor bolts:

[0132] The anchor bolts No. 1, No. 3, No. 5, No. 7, No. 9, No. 61, No. 63, No. 65, No. 73 and No. 75 on both sides of the roadway in front of the 21010 working face were selected, and the data from March to April were extracted and plotted into the anchor bolt force change curve ( Figure 12 ) for analysis.

[0133] The results show that: anchor rod No. 1 ( Figure 12(a) In March, as the working face advanced, the stress on the sidewall anchor rods remained stable at around 10 kN, indicating no significant deformation of the sidewall. However, in April, the stress on the sidewall anchor rods continued to rise, reaching a maximum of 83 kN on April 29, indicating an increasing trend in sidewall deformation. After the unloading intervention, the stress on the anchor rods dropped to around 20 kN. Anchor rod No. 3 ( Figure 5-12 (b)) and anchor rod No. 5 ( Figure 5-12 (c)) and anchor rod No. 9 ( Figure 5-12 The loads of anchor bolt (e)) have no significant changes and are stable at around 293kN, 13kN and 280kN respectively, which indicates that the surrounding rock structure of the tunnel is relatively intact. The data of anchor bolt No. 9 is missing after April 12. Figure 5-12 (d) The load was 97 kN at the beginning of March, and pressure relief measures were implemented on March 6 to reduce the load. However, as the working face advanced, the anchor load increased and eventually remained at around 108 kN, indicating that the roadway wall was bulging and that pressure relief measures needed to be implemented again.

[0134] Constructing a composite dynamic impact prevention and control system for thick hard rock coal pillars:

[0135] According to the cause analysis of the composite dynamic impact of the isolated working face during the mining process of the 21010 working face of the downhill coal pillar in the 21 mining area, it is found that as the working face advances, the overhang step of the overlying thick hard rock layer increases, causing the hard rock layer to suddenly break and release a large amount of energy, causing disturbance to the working face, thereby inducing impact danger. Therefore, for the overlying thick hard rock layer, in addition to considering the impact induced by the static load stress itself, the influence of the dynamic load stress must also be considered. Finally, based on this, this embodiment establishes a composite dynamic impact prevention and control system for the thick hard rock downhill coal pillar, such as Figure 13 shown.

[0136] This example determined the working face advance length to be 450 meters, ensuring it falls within the maximum fracture step distance of the thick hard rock strata and the protection of unmined coal pillars. Based on this, a comprehensive index method and a multi-factor coupling evaluation method were used to categorize the impact hazard areas of the 21010 working face, resulting in two weak and five moderate hazard areas.

[0137] Based on the hazard zone demarcation results, this example proposes impact prevention measures, primarily large-diameter drilling to relieve pressure on both sides, roof debonding, and floor debonding blasting to relieve pressure. Continuous monitoring is conducted during the prevention and control process, and large-diameter drilling to relieve pressure is increased in areas with abnormal monitoring.

[0138] This example verifies the effectiveness of the anti-bumping measures during the mining of the 21010 working face by comprehensively analyzing monitoring data such as microseismic, stress, and tunnel deformation. The microseismic monitoring results show that as the working face advances, the overhang step of the thick hard rock layer becomes larger, and the tiny cracks at its bottom release the accumulated energy, resulting in large-energy microseismic events. The number of large-energy microseismic events increased in April and May, but after the anti-bumping measures were implemented, the number of large-energy microseismic events in June decreased significantly. Mining stress analysis shows that online dynamic stress monitoring and staged unloading can control the stress value below the warning threshold. The stress analysis of the tunnel anchor rods shows that the tunnel wall went from a stage of no obvious deformation to a stage of obvious deformation, and then returned to a stage of no obvious deformation, which reflects the effectiveness of the dynamic unloading measures of the tunnel wall and successfully controls the deformation accumulation of the tunnel wall.

[0139] Finally, it should be noted that the above embodiments are only used to illustrate the technical solutions of the present invention and not to limit it. Although the present invention has been described in detail with reference to the above embodiments, ordinary technicians in the field should understand that the specific implementation methods of the present invention can still be modified or replaced by equivalents. Any modification or equivalent replacement that does not depart from the spirit and scope of the present invention should be covered by the scope of protection of the claims of the present invention.

Claims

1. A method for preventing and controlling the combined dynamic disaster of mine earthquake and rock burst caused by a huge and thick hard rock downhill coal pillar, characterized in that: include: Set up observation lines on the coal pillar working face to obtain surface settlement curves; Determine the advancement length of the coal pillar working face based on the surface settlement curve and numerical simulation results; The comprehensive index method is used to evaluate the impact hazard degree of the coal pillar working face and determine the rock burst hazard level; Dividing regions based on the rock burst hazard level; Conduct composite power impact monitoring in the divided areas and formulate composite power prevention and control measures.

2. The method according to claim 1, characterized in that The process of dividing the regions based on the rock burst hazard level includes: The multi-factor coupling evaluation method is used to analyze the impact factors of rock burst in the divided area. Based on the original rock deadweight stress field, the stress field superposition principle is used to synthesize the additional stress generated by tectonic stress, mining stress or other disaster-causing factors to obtain the total stress field strength of the coal rock. The rock burst hazard level areas are divided based on the ratio of the total stress field intensity to the uniaxial compressive strength.

3. The method according to claim 1, characterized in that The composite dynamic impact monitoring includes regional impact monitoring and local impact monitoring, and the local impact monitoring includes coal stress monitoring, roadway displacement monitoring, and support pressure monitoring; The regional impact monitoring process includes: arranging several groups of microseismic monitoring points along the direction of mining advancement, with the distance between adjacent monitoring points controlled within 200m to 300m, issuing an early warning when the daily cumulative energy detected exceeds 3.5E+05J; issuing an early warning when the energy release of a single vibration event reaches 2.0E+05J; and issuing an early warning when the density of high-frequency vibration events exceeds 5 times within 24 hours.

4. The method according to claim 3, characterized in that The coal body stress monitoring includes: arranging stress gauges in a double-hole gradient, with each group of stress gauges including two measuring points of 12m in a shallow hole and 18m in a deep hole, with the spacing of measuring points within the group along the roadway being ≤2m, and the spacing of adjacent monitoring groups being ≤30m. The stress gauges are arranged continuously along the roadway, with the first group no more than 40m away from the working face, and are installed on the east side of the transport chute, the west side of the return air chute, the east side of the auxiliary return air chute, and the upper side of the auxiliary cutting hole, at a distance of 0.5m to 1.5m from the roadway floor, and are moved forward with the working face mining; The tunnel displacement monitoring includes installing a set of roof separation and tunnel deformation monitoring equipment every 100m in the transport chute and return air chute; The support pressure monitoring includes installing a pressure monitoring station every three supports in the three-dimensional coal pillar area, and arranging them at a spacing of 10 supports in the extended coal pillar working face area. The periodic pressure situation of the working face is predicted based on the initial collapse step distance and periodic fracture of the direct roof and old roof. Before the working face is mined, the integrity index of the roof is counted and the activity pattern of the old roof is judged.

5. The method according to claim 1, characterized in that The composite power prevention and control measures include pre-pressure unloading measures during excavation and pressure unloading measures during mining. The pre-pressure unloading measures during excavation include pre-pressure unloading of large-diameter holes in the tunnel side during excavation and pre-pressure unloading of the broken bottom. The pressure unloading measures during mining include pre-pressure unloading measures of large-diameter holes in the tunnel side during mining, broken bottom measures and broken top measures.

6. The method according to claim 5, characterized in that The large-diameter drilling pre-pressure relief on the roadway side during excavation includes: drilling pressure relief holes perpendicular to the coal wall on both sides of the return air chute and the transport chute roadway; no pressure relief holes are drilled on the goaf side where the coal pillar is less than 15m; and pressure relief holes are drilled on the goaf side where the coal pillar is greater than 15m; Pressure relief holes are drilled on the upper side of the auxiliary cutting eye, but not on the lower side; Pressure relief holes are drilled on the east and west vertical coal walls of the auxiliary return air chute, and Φ 125mm pressure relief holes, with a spacing of no more than 3m and a hole depth of 15-18m; the spacing of pressure relief holes on the east side of the auxiliary return air chute shall be no more than 3m and a hole depth of 20-25m; Pressure relief holes are drilled on the upper side of the cut eye, but not on the lower side.

7. The method according to claim 5, characterized in that The pre-pressure relief of the broken bottom includes: taking floor pressure relief measures when the thickness of the bottom coal on one side of the roadway is greater than 1m; For thin low coal with thickness d satisfying 1m<d≤1.5m, pressure relief holes / water injection holes are arranged at the low angles of the double sides, and water is injected alternately through every other hole; For medium-thick low-lying coal with a thickness d of 1.5 < d ≤ 4m, water injection holes and blasting holes are constructed alternately at the low angles of the double sides, and high-pressure water injection is performed on the water injection holes; For thick low-lying coal with thickness d>4m, two bottom blasting holes are added every 10-20m on the basis of the pressure relief measures for medium-thick low-lying coal; No bottom blasting is carried out on the crosshead and the tunnel floor within 3m from the crosshead. Only bottom pressure relief holes and bottom water injection holes are constructed. The hole spacing is adjusted to 1m, and water is injected alternately through the holes.

8. The method according to claim 5, characterized in that The large-diameter drilling pressure relief measures for the roadway during mining include: On the basis of pre-unloading of the roadway side with large diameter drilling holes during the excavation, pressure relief holes are drilled at intervals of 1.5m in the high stress area of the cutting eye, and double-stage drilling is implemented on the east side within the 300m advance range of the transport drift advance area. For the auxiliary cutting eye and return air drift, inclination gradient control and spacing graded control are used to unload the heterogeneous coal body. For the return air drift, penetrating holes are drilled on the west side and the pressure relief range of the advance working face is not less than 150m.

9. The method according to claim 5, characterized in that The bottom breaking measures include: on the basis of the pre-unloading of the bottom, strengthening the monitoring of the bottom coal during the mining period, and continuing to take bottom breaking measures if the bottom coal still has the risk of impact. The bottom breaking parameters are based on the bottom breaking parameters during excavation, and are adjusted in time when the anti-impact requirements are not met.

10. The method according to claim 5, characterized in that The roof breaking measures include: drilling holes at intervals of 8m along the side roof of the coal pillar on the empty coal pillar greater than 10m in the working face transport chute and return air chute, and positioning the final hole at the mudstone-sandstone interbed interface.

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