Multi-disaster integrated prevention and control method for island working face in complex stress area

Abstract

The invention discloses a complex stress area island working face multi-disaster integrated prevention and control method which comprises the following steps: inducing fracture extension of a stress concentration area through hydraulic fracturing and deep hole blasting, weakening a top plate and reducing the area of a suspended roof; performing outburst prevention measure drilling on the coal seam construction completion area, and performing mechanical cave forming, pressure relief and anti-reflection drilling; the transportation drainage roadway is used for constructing crossing holes, pressure relief gas of an adjacent layer of a coal seam is extracted, and continuous pressure relief extraction is carried out on an unexploited coal seam of a bottom plate; and pressure relief gas extraction is conducted on adjacent layers, working faces are divided according to a high-risk area, a buffer area and a safe area according to the area verification condition in the extraction period, and zoning prevention and control are conducted for different areas. The method solves the bottleneck technology of island working disaster prevention and control in a high-stress concentration area under the disturbance of multiple remaining coal pillars, overcomes the problem of gas prevention and control of the island working face in a complex stress area, effectively relieves the main and essential contradiction of current coal mine safety production, and solves the problem that the actually measured drilling cutting desorption index K1 exceeds the standard during stoping.

Description

Technical Field

[0001] The present invention belongs to the technical field of coal mine disaster management, and specifically relates to an integrated multi-disaster prevention and control method for isolated island working faces in complex stress areas. Background Art

[0002] The final mining section of the working face is an isolated working face with multiple complex stress zones. During the mining process, due to stress concentration, intense mine pressure, and difficult ventilation and gas management, the roadway deformed severely and the mining efficiency was low. During the final mining period, a number of safety technical measures were implemented: construction hole drilling, hydraulic fracturing, deep hole blasting, mine pressure stress monitoring, supplementary drilling for face outburst prevention measures, and safe mining of the face. If effective measures are not taken in a timely manner, the desorption rate of coal seam gas will accelerate during the mining process under the influence of mining stress and the superimposed stress of multiple residual coal overburden, intensifying coal seam gas desorption and migration. Strong mine pressure phenomena such as dynamic loads are very likely to occur. Especially within the range of the working face's advanced support pressure, this may induce gas and roof disasters, seriously affecting production safety and causing economic losses and casualties. Summary of the Invention

[0003] In order to solve the above technical problems, the present invention provides an integrated multi-hazard prevention and control method for isolated island working faces in complex stress areas.

[0004] The technical solution adopted by the present invention is an integrated multi-hazard prevention and control method for isolated working faces in complex stress areas, which is:

[0005] The first step is to induce crack expansion in the stress concentration area through hydraulic fracturing and deep hole blasting, weaken the roof and reduce the overhanging roof area;

[0006] The second step is to drill holes for anti-burst measures in the coal seam construction area, and use mechanical drilling to create holes for pressure relief and permeability enhancement;

[0007] The third step is to drill holes through the layers in the transport and drainage tunnel to extract the gas from the adjacent layers of the coal seam and conduct continuous pressure relief extraction on the unmined coal seams on the floor.

[0008] Step 4: Divide the working face into high-risk areas, buffer zones and safe zones according to the regional verification results during mining, and carry out zoning prevention and control in different areas.

[0009] Furthermore, the hydraulic fracturing method is:

[0010] (1) Construct the fracturing borehole according to the design requirements, check the integrity of the borehole, observe whether the borehole has collapsed, and then clean the borehole to ensure that there is no obvious debris or sediment accumulation;

[0011] (2) Microseismic monitoring devices are placed in the return air lane and transport lane to monitor the initiation and expansion of hydraulic fracturing;

[0012] (3) Arrange the high-pressure pump and water tank at the designated location, connect the pipelines, install the pressure gauge, and conduct debugging; connect the water pipe to the tee and install the on / off valve control;

[0013] (4) Use a packer to seal the fracturing borehole, connect a high-pressure pump for hydraulic fracturing, and adopt single-hole multiple fracturing or single-hole single fracturing. The single fracturing time is 10 minutes from fracturing to initiation.

[0014] Furthermore, when drilling, record the location of mudstone or coal interlayer encountered, and avoid the location of mudstone or coal interlayer in the borehole during subsequent fracturing. During the pressurization process, the operation details are as follows: the three valves of the three-way valve are in the closed state, first adjust the water pump pressure to 5MPa, open the switch valve 1, observe the reading of the pressure gauge, and slowly open the switch valve 3, open it halfway first, wait for one minute and then fully open it, maintain the water pump pressure for 2 minutes, and adjust the water pump pressure to 5MPa, 10MPa, 15MPa, and 20MPa respectively. 20MPa, 25MPa, 30MPa, 35MPa; when the pressure is less than 20MPa, maintain the pump pressure for 2 minutes after each pressure adjustment. When it is greater than 20MPa, maintain the pump pressure for 1 minute until the pump pressure rises to 35MPa. At this time, the pressure gauge reading is 25MPa. Maintain the pump pressure for fracturing for 15 minutes. After fracturing, adjust the water pump pressure to 25MPa, then close the switch valve 3 to slowly release the pressure in the packer. After 10 minutes, open valve 2 to further release the pressure on the packer. When the pressure relief valve no longer discharges water, it means that the pressure relief is complete.

[0015] Furthermore, the deep hole blasting method is:

[0016] (1) Check the quality of the blasthole, requiring the hole position and angle to be accurate and meet the blasthole deflection requirements, and use φ50mmPVC to measure the actual depth of the drill hole and clean the rock powder residue in the drill hole;

[0017] (2) Load the explosives into the energy tube in advance and then put it into the drill hole. Cut off a long enough 2*1.0mm according to the length of the charge. 2 Cable, pass the leg wire through the energy-gathering tube and send it to the bottom of the hole along with the energy-gathering tube;

[0018] (3) A set of 4 sections of shaped tubes using φ63mm PVC are sent to the bottom of the hole. Two detonators are installed in the fifth section of shaped tubes. Repeat this step to push the remaining shaped tubes into the blasting position in the hole. Before the PVC tube filled with explosives is pushed into the drill hole and after it is pushed to the predetermined position, each conductor is subjected to a continuity test. Only after the continuity is qualified can the next operation be carried out.

[0019] (4) Use soft material to seal one end of the PVC tube that pushes the medicine roll, push the yellow mud into the mouth of the blast hole, use the joint to connect the next PVC tube, continue pushing until the yellow mud is pushed to the bottom of the medicine roll, pull the PVC tube to tamp the yellow mud, use the yellow mud to seal the hole, and use a bag and grouting to seal the hole.

[0020] Furthermore, there are two groups of drill holes for blasting, with each group having three holes in the return air tunnel and the transport tunnel; the explosives are installed using energy-gathering tubes, and when charging the drill holes, the first group of three holes are loaded first, and the second group is loaded after the blasting is completed, and only one group is blasted each time; the length of yellow mud placed in the drill holes each time shall not exceed 30 cm, the detonators in the drill holes are connected in parallel, and the wires between the drill holes are connected in series, and the personnel evacuation distance during detonation should not be less than 200 m from the blasting site, and it should not be on the same center line as the tunnel where the blasting site is located.

[0021] Furthermore, after the coal seam construction area anti-blowout measures drilling is completed and the effect is verified to be effective, mechanical hole drilling with a diameter of 500mm and a hole depth of 110m is constructed, with a drilling spacing of 5m.

[0022] Furthermore, conventional drilling, hydraulic fracturing, deep hole blasting, mechanical cavitation for pressure relief and permeability enhancement drilling, and advanced combined support measures are adopted in high-risk areas; conventional drilling and mechanical cavitation for pressure relief and permeability enhancement drilling are adopted in buffer areas; and conventional extraction drilling is adopted in safe areas.

[0023] Furthermore, the method also includes observing the mine pressure, including: observation of the working face support, observation of the activity pattern of the roof, observation of the tunnel deformation and observation of the working face support quality; the specific method is: ① A hydraulic support pressure gauge is installed on each hydraulic support of the comprehensive mining working face, and a measuring point is extracted every 10 supports. According to the pressure change of the column cylinder and the cyclic resistance increase, the initial collapse step and periodic activity pattern of the direct roof and the basic roof are judged, and the mine pressure manifestations such as the spalling and end face collapse of the coal wall of the working face are counted respectively to assist in judging the activity pattern of the basic roof and the direct roof; ② 5 points are selected for observation at the upper and lower ends and the advance support using a single column pressure gauge; ③ An observation station is set up every 50m for the surrounding rock of the upper and lower tunnels, and observations are made once a week, and records are made to analyze the roof pressure pattern; ④ The working face observation starts from the advancement of the mining working face to the first cycle of pressure, and each small shift observes once, and each round shift observes once after the cycle of pressure, and records are made.

[0024] Compared with the existing technology, the beneficial effect of the present invention is that, through the four-dimensional linkage coordinated prevention and control system of pressure relief-permeability enhancement-extraction-monitoring, hydraulic fracturing + mechanical cavitation + pressure relief and extraction of adjacent layers of the roof and floor plates + deep hole blasting + mine pressure monitoring and early warning are adopted to solve the bottleneck technology of disaster prevention and control of isolated island working in high stress concentration areas under multiple residual coal pillar disturbances, overcome the problem of gas prevention and control in isolated island working faces in complex stress areas, effectively alleviate the main contradictions in current coal mine safety production, and solve the problem of excessive K1 of drill cutting desorption index measured during mining. After the implementation of this technology, the regional verification K1 value during normal mining of the mining face is as low as 0.07mL / g·min 1 / 2 , maximum 0.17mL / g·min 1 / 2 The gas concentration of the return air flow did not exceed 0.30%, and the gas concentration in the upper corner did not exceed 0.38%, and the mining face was mined safely and smoothly. BRIEF DESCRIPTION OF THE DRAWINGS

[0025] Figure 1 It is a flowchart of the present invention;

[0026] Figure 2 It is the plan layout of the working surface and the adjacent relationship diagram;

[0027] Figure 3 It is the working face inclined profile layout and adjacent relationship diagram;

[0028] Figure 4 Layout of working face strike profile and adjacent relationships Figure 1-1 ;

[0029] Figure 5 Layout of working face strike profile and adjacent relationships Figure 2-2 ;

[0030] Figure 6 Borehole imaging for hydraulic fracturing drilling before fracturing;

[0031] Figure 7 Post-fracture drilling imaging for hydraulic fracturing drilling;

[0032] Figure 8 This is the relationship between fracturing pressure and time and hole depth for the first group of 90-meter boreholes in deep hole blasting;

[0033] Figure 9 This is a schematic diagram of the plan design of deep hole blasting in the working face;

[0034] Figure 10 This is the geological profile of the deep hole blasting of the working face Figure 1-1 ;

[0035] Figure 11 This is the geological profile of the deep hole blasting of the working face Figure 2-2 ;

[0036] Figure 12 This is the geological profile of the deep hole blasting of the working face Figure 3-3

[0037] Figure 13 This is a schematic diagram of the cross-section design of deep hole blasting in the working face;

[0038] Figure 14 This is a schematic diagram of the deep hole blasting results at the working face;

[0039] Figure 15 Schematic diagram of drilling design for creating holes in the working face;

[0040] Figure 16 Plan of drilling holes for decompression and gas extraction in the strata adjacent to the coal mining face;

[0041] Figure 17 Cross-section of the borehole for decompression and gas extraction in the strata adjacent to the coal mining face;

[0042] Figure 18 This is a plan diagram of the work face zoning prevention and control plan. DETAILED DESCRIPTION

[0043] The present invention will be further explained below with reference to the accompanying drawings to facilitate better understanding by those skilled in the art.

[0044] Example 1

[0045] like Figure 1-18 As shown, in order to achieve safe and efficient mining, an integrated multi-hazard prevention and control method for isolated island working faces in complex stress areas is proposed. The method is:

[0046] The first step is to induce crack expansion in stress concentration areas through hydraulic fracturing and deep hole blasting, weaken the roof, reduce the area of hanging roof, reduce the risk of gas accumulation in the upper corners, and reduce stress concentration in the coal seam.

[0047] The hydraulic fracturing construction method is:

[0048] (1) Construct the fracturing borehole according to the design requirements, check the integrity of the borehole, observe whether the borehole has collapsed, and then clean the borehole to ensure that there is no obvious accumulation of debris or mud and sand; the driller needs to record the location of the mudstone or coal interlayer encountered when drilling the hole, and avoid the mudstone or coal interlayer location in the borehole during subsequent fracturing.

[0049] (2) Microseismic monitoring devices are arranged in the return air channel and transport channel to monitor the initiation and expansion of hydraulic fracturing.

[0050] (3) Arrange the high-pressure pump and water tank at the designated location, connect the pipelines, install the pressure gauge, and conduct debugging; connect the water pipe to the tee and install a switch valve control.

[0051] (4) Use a packer to seal the fracturing borehole, and then connect a high-pressure pump for hydraulic fracturing. According to the design requirements, single-hole multiple fracturing or single-hole single fracturing is used, and the single fracturing time is 10 minutes after fracturing to initiation. The main process is: connect the packer → drill the hole to connect the water injection pipe → connect the high-pressure water pump → pressurize the packer → inject water for fracturing → stop the pump 10 minutes after water comes out of the adjacent hole or fracturing → release the pressure → start the second stage of fracturing → cyclic fracturing.

[0052] (5) During the pressurization process, the operation details are as follows: the three valves of the three-way valve are in the closed state. First, adjust the pump pressure to 5MPa, open the switch valve 1, observe the reading of the pressure gauge, and slowly open the switch valve 3, first halfway, wait for one minute and then fully open it, and maintain the pump pressure for 2 minutes. The water pump pressure is adjusted in sequence to 5MPa, 10MPa, 15MPa, 20MPa, 25MPa, 30MPa, and 35MPa; when it is less than 20MPa, maintain the pump pressure for 2 minutes after each pressure adjustment. When it is greater than 20MPa, maintain the pump pressure for 1 minute until the pump pressure rises to 35MPa. At this time, the pressure gauge reading is about 25MPa. Maintain the pump pressure for fracturing for 15 minutes. After fracturing, adjust the pump pressure to about 25MPa, then close the switch valve 3 and slowly release the pressure in the packer. After 10 minutes, open valve 2 to further release the pressure in the packer. When the pressure relief valve no longer discharges water, it means that the pressure relief is complete.

[0053] (6) Real-time monitoring during the fracturing process. If there is abnormal noise or large-scale water seepage on the roof, the fracturing will be stopped immediately. After the fracturing is completed, the next fracturing hole will be fractured.

[0054] Deep hole blasting:

[0055] First, check the quality of the blasthole. The hole position and angle must be accurate and meet the blasthole deflection requirements. Use φ50mm PVC to measure the actual depth of the drill hole and clean the rock powder residue in the drill hole. Due to the deep drill hole depth and large charge volume, in order to facilitate the filling of explosives and ensure the concentration of explosives, the explosives need to be pre-loaded into the energy-gathering tube before being loaded into the drill hole. Each energy-gathering tube is loaded with 2.4kg of explosives. Cut a long enough 2*1.0mm according to the charge length. 2 For the cable, insert the leg wire through the energy-gathering tube and feed it to the bottom of the hole. Insert a set of four φ63mm PVC energy-gathering tubes into the hole. Install two detonators in the fifth section. Repeat this process to push the remaining energy-gathering tubes into the hole to the blasting position. If there is excessive resistance during the pushing process, do not push forcefully. Instead, push slowly and gently back and forth. Do not pull on the wire. Before the PVC tube filled with explosives is pushed into the drill hole and after it is pushed to the desired position, conduct a continuity test on each wire. Only proceed to the next step if the continuity is satisfactory.

[0056] Use soft material to seal one end of the PVC tube that pushes the medicine roll, push the yellow mud into the blast hole, use the joint to connect the next PVC tube, and continue pushing until the yellow mud is pushed to the bottom of the medicine roll. Pull the PVC tube to tamp the yellow mud, use the yellow mud to seal the hole, and use a bag and grouting to seal the hole. To ensure the quality of the sealing, the length of the yellow mud inserted into the drill hole each time should not exceed 30cm, and the yellow mud sealed in the drill hole should be filled and tamped. During the sealing process, the copper core wire should be protected from damage and not brought into the blast hole to cause curling. After the sealing is completed, the wires should be tested for conductivity. Only after the conductivity is qualified can the next step be carried out.

[0057] To ensure simultaneous detonation of the explosive coils within the boreholes, the detonators within the boreholes are connected in parallel. To ensure simultaneous detonation of the boreholes with a single charge, the conductors between the boreholes are connected in series. Before connecting the busbar, conduct a continuity test on the busbar. Only proceed to the next step if the continuity is satisfactory.

[0058] Due to the large amount of explosives, personnel must evacuate at least 200m from the blasting site during detonation, and the evacuation distance should not be on the same center line as the tunnel where the blasting site is located. After reaching a safe distance, the conductor should be tested for continuity. Detonation can only be carried out after the continuity is qualified.

[0059] The second step is to drill holes in the coal seam construction area to prevent coal outbursts, and use construction machinery to create holes to relieve pressure and increase permeability.

[0060] After the coal seam construction area anti-blowout measures drilling is completed and the effect is verified to be effective, mechanical hole drilling with a diameter of 500mm and a hole depth of 110m is constructed, with a drilling spacing of 5m, to expand the coal body pressure relief range, increase the permeability of the coal seam, release local stress, increase gas extraction efficiency, and alleviate stress concentration.

[0061] The third step is to decompress gas extraction from adjacent layers: drill through the layers in the transport and extraction tunnel to extract decompress gas from the layers adjacent to the coal seam, and conduct continuous decompression extraction on the unmined coal seams on the bottom plate to block the gas migration path and reduce the gas influx into the working face.

[0062] Step 4: Based on regional verification during mining, the working face is divided into high-risk areas, buffer zones, and safe zones, and regional control measures are implemented for each area. Conventional drilling, hydraulic fracturing, deep hole blasting, mechanical cavitation, pressure relief and permeability enhancement drilling, and advanced combined support measures are used in high-risk areas; conventional drilling and mechanical cavitation, pressure relief and permeability enhancement drilling are used in buffer zones; and conventional extraction drilling is used in safe zones.

[0063] Furthermore, the mine pressure is observed at the same time. The mine pressure observation contents include: observation of working face supports, observation of roof activity patterns, observation of tunnel deformation and observation of working face support quality; the specific methods are: ① Install a hydraulic support pressure gauge on each hydraulic support of the fully mechanized mining working face, and extract one measuring point every 10 supports. According to the pressure change of the column cylinder and the cyclic resistance increase, the initial collapse step and periodic activity pattern of the direct roof and basic roof are judged, and the mine pressure manifestations such as spalling and end face collapse of the coal wall of the working face are counted respectively to assist in judging the activity pattern of the basic roof and direct roof; ② Select 5 points at the upper and lower ends and advance support and observe them using a single column pressure gauge; ③ Set up an observation station every 50m for the surrounding rock of the upper and lower tunnels, observe once a week, and keep records and analyze the roof pressure pattern; ④ Observe the working face from the start of the mining working face advancement to the first cycle of pressure, observe once for each small shift, and observe once for each round shift after the cycle of pressure, and keep records.

[0064] The final mining section of the 110305 working face at Songhe Company is an isolated working face with multiple complex stress zones. Located on the south wing of the 110304 working face and west of the 110306 goaf, the 110305 working face is 34 meters horizontally from the 110304 working face and 49 meters horizontally from the 110306 goaf. Unmined coal seams exist on the roof and floor. Influenced by the F113-9 fault, the 110305 working face has formed an isolated working face with complex stress zones. The 3# coal seam is 2.8 meters thick, has a dip of 26°, a mining length of 123 meters, and a strike length of 78 meters. The mining area of the final mining face at 110305 is 9,594 m. 2 During the mining process, due to stress concentration, severe mine pressure, and difficulty in ventilation and gas management, the tunnel deformation was serious (the amount of movement of the roof and floor was greater than 500mm / m); regional verification showed obvious signs of outburst; K1 approached the critical value (0.44mL / g·min 1 / 2 ); four gas warnings (maximum increase of 0.5%); an average daily mining recovery of 0.8 m, with low mining efficiency; and measured drill cuttings desorption index K1 exceeding the standard during mining at the 110305 working face. To achieve safe and efficient production, the aforementioned method was employed to implement integrated multi-hazard prevention and control measures for the isolated working face in the complex stress zone of the final mining section of the 110305 working face.

[0065] During hydraulic fracturing, the first set of boreholes, 90 meters deep and with a 36° inclination, were hydraulically fractured in the transport tunnel. These boreholes were fractured nine times, totaling 48 meters in length; the average maximum fracturing pressure was 22.88 MPa, and the average fracturing time was 23.7 minutes. During the fracturing process, water flowed from the fracturing boreholes to varying degrees, but the volume was relatively low. During the pressurization process, some sealed drill pipe connections leaked. After fracturing, valve 1 on the tee was closed, and the packer slowly depressurized, gradually reducing the pressure. When the pressure gauge on the tee dropped to 0, valve 2 was opened, and water gushed from the borehole. No water seepage occurred in the transport tunnel during the fracturing process.

[0066] Effect monitoring

[0067] Monitor the hydraulic support pressure in real time on site, compare it with the pressure before fracturing, and invert the roof collapse based on the hydraulic support pressure;

[0068] Monitor roof collapse using microseismic monitors located in the return air lane and transport lane;

[0069] Comparison before and after fracturing: The CXK7.2 mining borehole imaging instrument was used to observe and compare the hydraulic fracturing drilling before and after fracturing, and the fracturing effect was significant.

[0070] Blasting construction

[0071] Blasting location:

[0072] The first group: 20.5 meters after point C13 in the return air lane of 110305 and point B7 in the transport lane of 110305. Three holes were drilled in each of the return air lane and transport lane. The second group: 3 holes were drilled in each of the return air lane and transport lane at the stop line of the return air lane of 110305 and at the inclined transport lane of 110506.

[0073] The explosives are installed using shaped tubes. When charging the drill holes, first charge the first group of three holes, and then charge the second group after the blasting is completed; only one group is blasted each time.

[0074] The blasting construction parameters are shown in the following table:

[0075] Table 1 Parameters of deep hole blasting drilling in 110305 working face

[0076]

[0077]

[0078] Through-layer drilling was carried out in the 110906 transport and extraction tunnel to extract gas from the adjacent layers of coal seams 9, 6-2, 6-1, 5-2 and 5-1 to relieve pressure; continuous pressure relief extraction was carried out on the unmined coal seams on the bottom plate to block the gas migration path and reduce the influx of gas into the working face.

[0079] The stress online monitoring system monitors stress distribution and identifies abnormal coupling signals between stress and gas. Analyze pressure manifestation monitoring data to understand the deformation and stress variation patterns of the roadway surrounding rock. This allows for the development of roadway pressure manifestation patterns during mining, predicts the development trend of roadway pressure manifestation, and evaluates roadway support effectiveness. This allows for the immediate understanding of underground mining pressure manifestation patterns, providing a basis for optimizing roadway support parameters.

[0080] Effect evaluation

[0081] index Before governance After governance Maximum K1 value 0.42mL / g·min1 / 2 0.17mL / g·min1 / 2 Return air flow gas 0.4％-0.5％ Stable ≤0.30% Upper corner gas Peak 1.2% ≤0.38% (continuous monitoring) Ergonomics The average mining rate is 0.8m / d The overall advancement speed of the mining face has increased by 3 times

[0082] The 110305 working face was mined normally during the last mining period. The drilling main of the 110906 drainage lane was stable at more than 45%, and the pure extraction volume was 16m 3 / min; after taking measures, the regional verification K1 value was as follows: minimum 0.07mL / g·min1 / 2, maximum 0.17mL / g·min1 / 2, return air flow did not exceed 0.30%, and upper corner gas did not exceed 0.38%.

[0083] Songhe Coal Mine adopts a four-dimensional coordinated prevention and control system of pressure relief-permeability enhancement-extraction-monitoring to overcome the problem of gas prevention and control in isolated working faces in complex stress areas, effectively alleviate the main contradictions in current production safety, avoid production stoppage losses caused by disasters, and recover economic losses based on the average daily production capacity, indirectly saving 4.5 million yuan, achieving normal mining of the 110305 working face of Songhe Company, and safely producing about 35,000 tons of coal in high-risk areas and buffer zones within the complex stress zone. The clean coal recovery rate of the 3# coal seam is 55%, and calculated at 1,330 yuan / ton, the economic benefits are 25.6 million yuan.

[0084] The above-described embodiments are merely descriptions of preferred embodiments of the present invention and are not intended to limit the scope of the present invention. Any modifications and improvements made to the technical solution of the present invention by those skilled in the art without departing from the spirit and principles of the present invention shall fall within the scope of protection defined by the claims of the present invention.

Claims

1. An integrated multi-hazard prevention and control method for isolated working faces in complex stress areas, characterized by: The method is, The first step is to induce crack expansion in the stress concentration area through hydraulic fracturing and deep hole blasting, weaken the roof and reduce the hanging roof area; The second step is to drill holes for anti-outburst measures in the coal seam construction area, and use mechanical drilling to create holes for pressure relief and permeability enhancement; The third step is to drill holes through the layers in the transport and drainage tunnel to extract the gas from the adjacent layers of the coal seam and conduct continuous pressure relief extraction on the unmined coal seams on the floor. The fourth step is to depressurize the adjacent layers and extract the gas. According to the regional verification during the mining period, the working face is divided into high-risk areas, buffer zones and safe areas, and zoning prevention and control are carried out for different areas.

2. The multi-hazard integrated prevention and control method for isolated working faces in complex stress areas according to claim 1 is characterized in that: The hydraulic fracturing method is: Construct the fracturing borehole according to the design requirements, check the integrity of the borehole, observe whether there is any collapse in the borehole, and then clean the borehole to ensure there is no obvious debris or sediment accumulation; Microseismic monitoring devices are installed in the return air lane and transport lane to monitor the initiation and expansion of hydraulic fracturing; Place the high-pressure pump and water tank at the designated location, connect the pipelines, install the pressure gauge, and conduct commissioning; connect the water pipe to the tee and install an on / off valve control; Use a packer to seal the fracturing borehole, connect a high-pressure pump for hydraulic fracturing, and adopt single-hole multiple fracturing or single-hole single fracturing. The single fracturing time is 10 minutes from fracturing to initiation.

3. The multi-hazard integrated prevention and control method for isolated working faces in complex stress areas according to claim 2 is characterized in that: When drilling, record the location of mudstone or coal interlayer encountered, and avoid the location of mudstone or coal interlayer in the borehole during subsequent fracturing. During the pressurization process, the operation details are as follows: the three valves of the three-way valve are in the closed state, first adjust the pump pressure to 5MPa, open the switch valve 1, observe the reading of the pressure gauge, and slowly open the switch valve 3, open it halfway first, wait for one minute and then open it completely, maintain the water pump pressure for 2 minutes, and adjust the water pump pressure to 5MPa, 10MPa, 15MPa, 20MPa respectively. Pa, 25MPa, 30MPa, 35MPa; when it is less than 20MPa, maintain the pump pressure for 2 minutes after each pressure adjustment. When it is greater than 20MPa, maintain the pump pressure for 1 minute until the pump pressure rises to 35MPa. At this time, the pressure gauge reading is 25MPa. Maintain the pump pressure for fracturing for 15 minutes. After fracturing, adjust the water pump pressure to 25MPa, then close the switch valve 3 to slowly release the pressure in the packer. After 10 minutes, open valve 2 to further release the pressure on the packer. The pressure relief valve no longer discharges water, indicating that the pressure relief is complete.

4. The multi-hazard integrated prevention and control method for isolated working faces in complex stress areas according to claim 1 is characterized in that: The deep hole blasting method is: Check the quality of blastholes, ensure the hole location and angle are accurate, meet the blasthole deflection requirements, measure the actual depth of the borehole, and clean the rock dust and slag in the borehole; Pre-load the explosives into the energy-gathering tube and then load it into the drill hole. Cut a 2*1.0mm² cable of sufficient length according to the length of the explosive charge, insert the leg wire through the energy-gathering tube, and feed it to the bottom of the hole along with the energy-gathering tube. Push a set of 4-section focused tubes into the bottom of the hole, install 2 detonators in the 5th section, and repeat this process to push the remaining focused tubes into the blasting position in the hole; Before the PVC pipe filled with medicine rolls is pushed into the drill hole and after it is pushed to the predetermined position, each conductor is subjected to a continuity test. Only after the continuity is qualified can the next operation be carried out; Use soft material to seal one end of the PVC tube that pushes the medicine roll, push the yellow mud into the mouth of the blast hole, use the joint to connect the next PVC tube, continue pushing until the yellow mud is pushed to the bottom of the medicine roll, pull the PVC tube to tamp the yellow mud, use the yellow mud to seal the hole, and use a bag and grouting to seal the hole.

5. The multi-hazard integrated prevention and control method for isolated working faces in complex stress areas according to claim 4 is characterized in that: There are two groups of drill holes for blasting, with three holes in each group in the return air tunnel and the transport tunnel; the explosives are installed using energy-gathering tubes. When charging the holes, the first group of three holes are loaded first, and the second group is loaded after the blasting is completed. Only one group is blasted each time; the length of yellow mud placed in the holes shall not exceed 30cm each time, the detonators in the holes are connected in parallel, and the wires between the holes are connected in series. When detonating, the evacuation distance of personnel should not be less than 200m from the blasting site, and it should not be on the same center line as the tunnel where the blasting site is located.

6. The multi-hazard integrated prevention and control method for isolated working faces in complex stress areas according to claim 1 is characterized in that: After the coal seam construction area anti-blowout measures drilling is completed and the effect is verified to be effective, mechanical hole drilling with a diameter of 500mm and a hole depth of 110m is constructed, and the drilling spacing is 5m.

7. The multi-hazard integrated prevention and control method for isolated working faces in complex stress areas according to claim 1 is characterized in that: For high-risk areas, conventional drilling, hydraulic fracturing, deep hole blasting, mechanical cavitation, pressure relief and permeability enhancement drilling, and advanced combined support measures are adopted; for buffer zones, conventional drilling and mechanical cavitation, pressure relief and permeability enhancement drilling are adopted; and for safe zones, conventional extraction drilling is adopted.

8. The multi-hazard integrated prevention and control method for isolated working faces in complex stress areas according to claim 1 is characterized in that: The method also includes observing the mine pressure, including: observation of the working face support, observation of the activity pattern of the roof, observation of the tunnel deformation and observation of the working face support quality; the specific method is: ① Install a hydraulic support pressure gauge on each hydraulic support of the fully mechanized working face, and extract one measuring point every 10 supports. According to the pressure change of the column cylinder and the cyclic resistance increase, the initial collapse step and periodic activity pattern of the direct roof and basic roof are judged, and the mine pressure manifestations such as the spalling and end face collapse of the coal wall of the working face are counted respectively to assist in judging the activity pattern of the basic roof and the direct roof; ② Select 5 points at the upper and lower ends and the advance support and observe them using a single column pressure gauge; ③ Set up an observation station every 50m for the surrounding rock of the upper and lower tunnels, observe once a week, and keep records and analyze the roof pressure pattern; ④ Observe the working face from the start of the mining working face advancement to the first cycle of pressure, observe once for each small shift, and observe once for each round shift after the cycle of pressure, and keep records.

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