A roof blasting pressure relief method based on stratum division and weighted comprehensive score

Abstract

This invention discloses a roof blasting pressure relief method based on stratigraphic division and weighted comprehensive scoring, belonging to the field of roof blasting pressure relief. The method includes: obtaining geological condition parameters of the working face under test through field surveys; dividing the main fault and secondary caving layers and calculating the theoretical caving zone stratum thickness; determining blasting parameters, including determining the roof-breaking blasting height, borehole layout, blasting angle, borehole depth, and single-hole charge; deploying microseismic sensors and conducting downhole monitoring; and evaluating the roof-breaking effect after blasting. This method can quickly set blasting parameters for roof-breaking technology and rapidly evaluate the blasting effect to adjust subsequent blasting operations. It significantly improves the safety and economy of roof-breaking blasting, is suitable for deep mines, complex geological conditions such as composite roofs, and provides a systematic solution for rockburst prevention.

Description

Technical Field

[0001] This invention relates to the field of roof blasting and pressure relief, specifically to a roof blasting and pressure relief method based on stratigraphic division and weighted comprehensive scoring. Background Technology

[0002] Thick, hard rock strata overhanging a large area within the caving zone above a coal seam can release enormous energy upon fracturing, potentially triggering rockbursts. Therefore, roof blasting for pressure relief in thick, hard rock strata is an effective means of preventing rockbursts. The main objectives of roof blasting for pressure relief are: reducing the length of the overhanging roof, decreasing the intensity of periodic pressure, and reducing the probability of high-energy mine tremor events. Existing technologies and patents have achieved these objectives to some extent, but the following problems remain: ① Poor blasting effect: The lack of a systematic division of the rock strata above the coal seam makes it impossible to accurately locate the target rock strata for blasting; ② Lack of basis for blasting parameter design: Key parameters such as borehole depth and borehole inclination angle are determined based on experience, resulting in a high failure rate; ③ Lack of an effect evaluation system: There is a lack of scientific and objective methods to evaluate the blasting effect, a lack of treatment measures based on the evaluation results, and an incomplete process system. Therefore, there is an urgent need for a technical method and effect evaluation method for roof blasting in coal mines to solve the above problems. Summary of the Invention

[0003] To address the problems existing in the prior art, this invention provides a roof blasting decompression method based on stratigraphic division and weighted comprehensive scoring.

[0004] The present invention adopts the following technical solution: A roof blasting decompression method based on stratigraphic division and weighted comprehensive scoring includes the following steps: Step 1: Conduct on-site investigation to obtain geological condition parameters of the working face to be tested; Step 2: Divide the main fault and the secondary collapse zone, and calculate the theoretical thickness of the strata in the collapse zone; Step 3: Determine the blasting parameters, including determining the blasting height of the roof break, the arrangement of the blast holes, the blasting angle, the blast hole depth, and the charge per hole; Step 4: Deployment of microseismic sensors and downhole monitoring; Step 5: Evaluate the effect of the blasting to break the roof after the blast.

[0005] Preferably, step 2 specifically includes: There are several rock strata above the coal seam, which are, from bottom to top: the nth stratum, the (n+1)th stratum, ... the (n+i-1)th stratum, the (n+i)th stratum, and the (n+i+1)th stratum; Main fault layer division: The load applied by the (n+i+1)th rock layer to the nth rock layer ( W n+i+1 ) n The load exerted on the nth rock layer by the (n+i)th rock layer is less than that exerted on the nth rock layer.W n+i ) n Then the nth layer is the main fault layer; Sub-collapse layer classification: Layers from n+1 to n+i are considered sub-collapse layers; The load is calculated as follows:

[0006] In the formula, W n The load of the nth layer itself is kN; The load of the nth rock layer on the 1st rock layer is expressed in kN. E The elastic modulus is expressed in MPa. The thickness of the nth rock layer; Let n be the unit weight of the nth rock stratum; The formula for calculating the theoretical caving zone thickness is:

[0007] In the formula, D The effective mining height is in meters (m). Σh The theoretical thickness of the caving zone is in meters (m). This represents the average fracture expansion coefficient of the rock strata in the caving zone.

[0008] Preferably, step 3 specifically includes: The method for determining the blasting height of the fault is as follows: blast the main fault strata within the theoretical caving zone thickness range. That is, by determining the number and location of the main faults within the theoretical caving zone thickness, blasting is carried out at the height of the main fault locations. This method is divided into the following two types: (1) When the main fault layer n is determined, the height of the rock layer is equal to or exceeds the theoretical thickness of the collapsed rock layer. Σ h In this case, the upper surface of the nth main fault layer is the highest point of the blasting height; (2) If the number and thickness of the main fault layers are too high and have not yet reached the theoretical thickness of the collapsed rock layers, then the fault is considered to be faulty. Σh The economic efficiency of blasting is too low. Taking into account the impact of main faults of different heights on mine pressure, the main faults with rock strata height within 30-50m above the coal seam are first blasted to break the top. Then, rockburst is prevented by a combination of measures, namely, pressure relief through coal seam drilling and strengthening of active and passive support of roadways. Arrangement of blast holes: The blast holes are arranged in groups. In solid coal roadways, there are 3 blast holes in each group; in goaf roadways, there are 4 blast holes in each group. The borehole diameter is 80+ mm and the group spacing is 15 m. The roof blasting holes are started 40 m away from the cut-out of the two roadways and continue to a position 50 m away from the stop line. Among them, the borehole angle of borehole No. 1 in the solid coal roadway The angle of borehole #2 The angle of borehole #3 It is 60°; Calculate according to the following formula : ; ; In the formula: The angle between the borehole axis and the horizontal plane; The actual height of the target main fault. The working face width is the length of the working face that advances along the dip of the coal seam. The distance between the tops of boreholes #1 and #2; The angle of the blast hole at blast hole #4 along the tunnel The borehole angle for borehole #5 is 40°~60°. The borehole angle for borehole #6 is 60°~70°. The borehole angle is 60° for borehole #7. It is 50°; The depth of each borehole was calculated based on geometric relationships. : ; m takes the integer value from 1 to 7; Calculate the charge amount per orifice using the following formula. : .

[0009] In the formula: Explosive density; For the length of the charge section, take... ; This refers to the cross-sectional area of ​​the explosive.

[0010] Preferably, the microseismic sensor arrangement is as follows: no less than 6 microseismic sensors are arranged within a 1500m radius around the working face, with two microseismic sensors in the same tunnel being 400-450m apart. Microseismic sensors closer to the working face are moved when they are less than 100m away from the working face. The positions of the microseismic sensors are adjusted in a timely manner according to the actual site conditions to meet the microseismic monitoring requirements. Microseismic event monitoring: Using a microseismic monitoring system, the measuring points are monitored in real time during the mining operation, and the seismic source is located and the microseismic energy is calculated every day.

[0011] Preferably, step 5 specifically includes: The peak energy value 'a' of the microseismic events after the blasting roof break was obtained, and the number of microseismic events 'b' of the blasting roof break was obtained. The peak working resistance c of the working face support after the blasting roof break is obtained, and the periodic pressure step d after the blasting roof break is obtained; Normalization is performed: ; ; ; ; In the formula, This represents the maximum energy value of the microseismic event. This represents the average energy of the microseismic events. This represents the minimum energy value of the microseismic event. This represents the maximum number of microseismic events. This represents the average number of microseismic events. This represents the minimum number of microseismic events. This represents the maximum working resistance of the support. This represents the average working resistance of the support. This represents the minimum working resistance of the support. The maximum step size is determined by the period; The average step size is used to control the periodicity. The minimum step size is determined by the periodicity. As the peak energy weight of microseismic events, As the weight of the number of microseismic events, As the peak resistance weight of the stent, The peak resistance of the stent is weighted. Evaluation score before blasting the roof: S 前 = ×ω a1 + ×ω b1 + ×ω c1 + ×ω d1 ; Evaluation score after blasting and breaking the roof: S 前 = ×ω a2 + ×ω b2 + ×ω c2 + ×ω d2 ; In the formula, This represents the average energy of micro-seismic events prior to the blasting and roof collapse. The average value of microseismic energy after blasting and roof breaking. This represents the average number of microseismic events prior to the blasting roof break. This represents the average number of microseismic events following the blasting and roof break. This represents the average working resistance of the support before the blasting roof break; This represents the average working resistance of the support after the blasting and roof break. The average step distance for periodic pressure application before blasting to break the roof; This represents the average step distance for periodic pressure application after blasting to break the roof. The weight of the peak energy of microseismic events before the blasting fracture is given. Weighting of the number of microseismic events prior to the blasting failure. The peak resistance of the support before the blasting roof break is weighted. The peak resistance of the support before the blasting roof break is weighted. The weight of the peak microseismic energy after blasting and roof breaking. Weighting of the number of microseismic events following blasting and roof collapse. The peak resistance of the support structure after blasting and roof breakage is weighted. The peak resistance of the support after blasting and roof breakage is weighted. Judgment formula: S = (S 后 / S 前 ) × 100%; S represents the proportion of residual energy in the roof after blasting. The smaller the S value, the better the blasting effect. When S≤60%, the blasting effect is excellent; when 60%≤S≤80%, the blasting effect is good; and when S≥80%, the blasting effect is poor.

[0012] The beneficial effects of this invention are: (1) This invention is the first to theoretically calculate and divide the rock strata above the coal seam into the "main fault" and the "secondary caving layer", and adjust the blasting height and range in combination with the theoretical thickness of the caving zone. This technology solves the problem that traditional methods cannot accurately locate the target rock strata, greatly improves the blasting effect and reduces the overhang length behind the working face.

[0013] (2) This invention provides a dynamic design method for blasting parameters and a full-process construction technology based on the identification of the main fault. It innovatively proposes a three-dimensional parameter matching mechanism of "tunnel type-advance distance-coverage range". Compared with the traditional empirical design method, this method is more scientific and reliable. It also develops an underground adaptive charging and filling device, which can achieve high-precision positioning of charging while reducing the intensity of manual operation, effectively avoid the explosive getting damp and energy leakage, and significantly improve the safety and economy of blasting.

[0014] (3) This invention constructs a dynamic evaluation system for blasting effects based on a multi-index weighted comprehensive scoring method. It innovatively introduces a four-dimensional evaluation model consisting of "micro-seismic energy peak (a) - number of events (b) - peak support resistance (c) - periodic pressure step distance (d)". The evaluation effect is quantified through normalization formulas and roof failure judgment methods, and is accompanied by a closed-loop management strategy of "evaluation-early warning-control". It significantly improves the safety and economy of roof failure blasting, and is suitable for complex geological conditions such as deep mines and composite roofs, providing a systematic solution for rockburst prevention and control. Attached Figure Description

[0015] Figure 1 This is a flowchart of the present invention.

[0016] Figure 2 This is the division of the main fault layer and secondary collapse layer in the upper layer of the mine roof in Example 1.

[0017] Figure 3 This is a schematic diagram of a composite beam.

[0018] Figure 4 A schematic diagram of the layout of blast holes in solid coal roadways and goaf roadways.

[0019] Figure 5 for Figure 4 Schematic diagram of the layout and orientation of boreholes 1, 2, 4, and 5 in the cross-section I-I solid coal roadway and goaf roadway.

[0020] Figure 6 for Figure 4 profile Figure II-II Schematic diagram of the layout of borehole #3 in the solid coal roadway.

[0021] Figure 7 for Figure 4 profile Figure III-III A schematic diagram of the layout of blast holes #6 and #7 along the coal seam roadway.

[0022] Figure 8 This is a schematic diagram of a new roof pre-splitting blasting charge filling device.

[0023] Figure 9 for Figure 8 Enlarged view of a specific area.

[0024] Figure 10 This is a schematic diagram of the arrangement of micro-vibration sensors on the working face.

[0025] Figure 11 This is a design drawing for the axial partitioning of the borehole. Detailed Implementation

[0026] The specific embodiments of the present invention will be further described below with reference to the accompanying drawings and specific examples: Combination Figures 1 to 11 A roof blasting decompression method based on stratigraphic division and weighted comprehensive scoring includes the following steps: Step 1: Conduct on-site investigation to obtain geological condition parameters of the working face to be tested.

[0027] Specifically, this includes: (1) determining the geological structure of the mine: the classification of geological types and the properties of coal-bearing strata; (2) Previous mining earthquake occurrences: frequency, location, and energy released in previous mining earthquakes; (3) Determine the impact tendency of the relevant working face: impact tendency of the working face and roadway; (4) Changes in working face support pressure: periodic pressure step distance and peak support pressure.

[0028] Step 2: Divide the main fault layer and the secondary collapse layer, and calculate the thickness of the rock strata in the theoretical collapse zone.

[0029] Theoretical calculations were used to identify the rock strata above the coal seam that play a decisive role in the occurrence of rockbursts.

[0030] In the roof strata above a coal seam, several adjacent strata deform synchronously to form a composite beam. The thick, hard rock layer below controls all synchronously deforming strata within the composite beam; this thick, hard rock layer is called the "primary fault" of the composite beam, and the other strata within the composite beam are called "secondary caving layers." For example... Figure 3 .

[0031] To simplify the calculation, we assume that the load on each rock layer is uniformly distributed, and calculate the load on each rock layer itself and the load exerted on it by the rock layer above.

[0032] Above the coal seam are several rock strata, from bottom to top: the nth stratum, the (n+1)th stratum, ..., the (n+i-1)th stratum, the (n+i)th stratum, and the (n+i+1)th stratum; n and i are both integers; n ≥ 0, i > 1.

[0033] Main fault layer division: The load applied by the (n+i+1)th rock layer to the nth rock layer ( W n+i+1 ) n The load exerted on the nth rock layer by the (n+i)th rock layer is less than that exerted on the nth rock layer. W n+i ) n If the nth layer is the main fault layer, then the nth layer is the main fault layer.

[0034] Sub-collapse layer definition: Layers from n+1 onwards (excluding layer n) up to layer n+i (including layer n+i) are considered sub-collapse layers.

[0035] The load calculation method is as follows: .

[0036] In the formula,W n The load of the nth floor itself is kN; W n ) 1 The load of the nth rock layer on the 1st rock layer is expressed in kN. E The elastic modulus is expressed in MPa. The thickness of the nth rock layer; Let n be the unit weight of the nth rock stratum.

[0037] Theoretically, after some rock strata collapse and fragment, they can fill the free space of the goaf; the thickness of this portion of rock strata is the thickness of the collapse zone. To determine the theoretical thickness of the collapse zone above the coal seam, it is necessary to consider the effective mining height, the theoretical collapse zone thickness, and the average fragmentation coefficient of the collapse zone strata.

[0038] The formula for calculating the theoretical caving zone thickness is: .

[0039] In the formula, D The effective mining height is in meters (m). Σh The theoretical thickness of the caving zone is in meters (m). This represents the average fracture expansion coefficient of the rock strata in the caving zone.

[0040] Step 3: Determine the blasting parameters, including determining the blasting height of the roof break, the arrangement of the blast holes, the blasting angle, the blast hole depth, and the charge per hole.

[0041] The purpose of roof blasting for pressure relief is to place explosives at the height of the target rock layer (the main fault). The blasting causes the target rock layer to be filled with fissures. When the coal seam is mined to the area below the target rock layer, it collapses in time to fill the goaf. This prevents rockburst accidents caused by large-area overhang and sudden fracture of thick, hard rock layers releasing energy.

[0042] The method for determining the blasting height of the fault is as follows: blast the main fault strata within the theoretical caving zone thickness range. That is, by determining the number and location of the main faults within the theoretical caving zone thickness, blasting is carried out at the height of the main fault locations. This method is divided into the following two types: (1) When the main fault layer n is determined, the height of the rock layer (including the main fault layer n and its associated secondary collapse layers, and the rock layer below the main fault layer n) is equal to or exceeds the theoretical collapse rock layer thickness. Σh In this case, the upper surface of the nth main fault layer is the point with the highest blasting height.

[0043] (2) If the number and thickness of the main fault layers are too high and have not yet reached the theoretical thickness of the collapsed rock layers, then the fault is considered to be faulty. ΣhThe economics of blasting are too low. Taking into account the impact of main faults at different heights on mine pressure, the main faults with rock strata height within 30-50m above the coal seam are first blasted to break the top. Then, rockburst is prevented by a combination of measures, namely, pressure relief through coal seam drilling and strengthening of active and passive roadway support.

[0044] The main influencing factors in determining the borehole layout, blasting angle, and borehole depth are: (1) Advance distance: Advance roof blasting must be carried out at a certain distance from the working face, away from the mining activity area of ​​the working face, to reduce the disturbance of blasting vibration to the coal body and the roof of the goaf.

[0045] (2) Coverage area: ① The total length of the blast hole is divided into a charging section and a sealing section. To ensure the blasting effect (prevent explosive leakage, allow energy to accumulate in the blast hole, and ensure operational safety), the length of the sealing section shall not be less than 1 / 3 of the total length of the blast hole.

[0046] ② The design of the borehole depth and blasting angle needs to be calculated: by adjusting the borehole depth and angle, the blasting can cover the main fault within a range of 30-50m above the coal seam; there should be a certain interval between the tops of two adjacent boreholes (generally not less than 10m).

[0047] (3) Roadway type: Solid coal roadways and roadways along the goaf are different in terms of mining stress, roof structure, impact risk, borehole depth and blasting angle arrangement.

[0048] Arrangement of blast holes: The blast holes are arranged in groups. In solid coal roadways, there are 3 blast holes in each group; in goaf roadways, there are 4 blast holes in each group. The borehole diameter is 80+ mm and the group spacing is 15 m. The roof blasting holes are constructed starting 40 m away from the cut-out of the two roadways and continuing to a position 50 m away from the stop line.

[0049] Requirements for borehole layout: (1) Solid coal roadway: ① Advance distance requirement: During the mining process, the blasting and breaking holes in the solid coal roadway must always be at least 300m ahead of the working face.

[0050] ② Hole depth and blasting angle: Each group has 3 blast holes. The azimuth angle of blast hole #1 and blast hole #2 is 90°. and borehole angle The borehole depth is determined precisely based on the location of the main fault, ranging from 40° to 50° and 50° to 60°, as shown in the formula below. The borehole depth also needs to be precisely determined based on the location of the main fault. The azimuth of borehole #3 is 180°, and the borehole angle... The angle is 60°, and the hole depth needs to be precisely determined based on the location of the main fault.

[0051] Calculate according to the following formula : ; .

[0052] In the formula, The actual height of the target main fault. The working face width is the length of the working face that advances along the dip of the coal seam. The distance between the tops of the two blast holes, #1 and #2.

[0053] (2) Along the tunnel: ① Advance distance requirement: During the mining process, the blasting hole along the goaf should always be at least 350m ahead of the working face.

[0054] ② Hole depth and blasting angle: Four blasting boreholes are arranged in each group along the tunnel. The azimuth angle of borehole #4 and borehole #5 is 270°. The blasting angle of borehole #4 is... The borehole angle for borehole #5 is 40°~60°. The azimuth angle is 60°~70°, and the hole depth needs to be precisely determined based on the location of the main fault. The azimuth angle of boreholes #6 and #7 is 180°. The borehole angle of borehole #6... The borehole angle is 60° for borehole #7. The angle is 50°, and the borehole depth needs to be precisely determined based on the location of the main fault. For example... Figure 4-7 As shown.

[0055] The depth of each borehole was calculated based on geometric relationships. : ; m can be any integer from 1 to 7. The depth of borehole #1. This represents the depth of borehole #2, and so on.

[0056] Calculate the charge amount per orifice using the following formula. :

[0057] In the formula: Explosive density; For the length of the charge section, take... ; This refers to the cross-sectional area of ​​the explosive.

[0058] At the same time, a new "roof pre-splitting explosive filling device" was proposed, such as... Figure 8 and Figure 9 This device significantly reduces the manual labor intensity of blasting and improves the success rate of blasting: (1) The device has the following advantages: ①Automatic loading: Guided by the guide plate 6, automatic lifting device 5, and splicing sleeve 4, the explosive can be accurately delivered to the designated location. Compared with the previous delivery method, this improves loading efficiency, reduces labor, and the cylindrical structure protects the explosive from moisture and failure due to the humid underground environment.

[0059] ② "Delivery + Filling + Fixing" integrated structure: The pushing structure delivers the explosive to the designated position, and the spraying device 8 sprays and fills the gap between the explosive and the rock layer, reducing the air cavity and stabilizing the blasting explosive in the designated position while improving the effect of blasting to break the roof.

[0060] ③ It has the function of "segmented loading": multiple segments of explosives are loaded into the device. The pushing structure transports the first segment of explosives to the designated position (deepest point), and the spraying device 8 can immediately fill and fix the first segment of explosives. The pushing structure continues to transport the second segment of explosives to the designated position, and the spraying device can fill and fix the second segment of explosives again. In this way, the device can realize segmented loading of explosives at multiple designated positions.

[0061] ④ Prefabricated filling materials: Different filling materials can be prefabricated for tunnels with different rock types, ensuring full coupling between the explosives and the rock strata. Furthermore, the filling materials can be prefabricated on the surface in one go, eliminating the need for on-site mixing underground and reducing blasting costs.

[0062] ⑤ Explosion-proof equipment: The motor used is a special underground mining safety motor with explosion-proof and water-spray-proof functions. The motor fan and bearings are coated with anti-static material to prevent the generation of sparks. It is suitable for harsh environments such as high gas and humid water.

[0063] (2) The device comprises the following structures: ① The pushing structure includes a support base 1, a support frame 2, a pushing motor 3, a splicing sleeve 4, an automatic lifting device 5, and a pushing guide plate 6. The support base is connected to the support frame, the support frame has a pushing motor on the upper part, the pushing motor is spliced ​​with the automatic lifting machine, the automatic lifting machine has a guide plate on the upper part, and is spliced ​​with the splicing sleeve. With the guiding action of the guide plate and the splicing sleeve, and driven by the pushing motor, the automatic lifting machine sends the explosive into the hole through the guide plate.

[0064] ② Spraying structure: including a conveying pipe 7 in the splicing sleeve, a spraying device 8 at the pipe end, an interface 9 at the bottom of the pipe, a control valve 10 at the interface, a replaceable filling device at the connection of the control valve, and the control valve pressurizes and conveys the filling material through a motor. During the spraying process, the replaceable spraying device is first connected to the lower interface. After the power is turned on 11, the control valve pressurizes the device so that the drug can reach the end of the pipeline.

[0065] ③ Anti-friction structure: including the sliding plate 12 of the inner wall of the splicing sleeve and the lubricant. During the pushing process, the sliding plate and the lubricant prevent the inner wall from rubbing against the explosive.

[0066] ④ Moisture-proof and fire-proof mechanism: including motor, the outer shell is made of stainless steel or high-strength aluminum alloy, the surface is sprayed with epoxy resin anti-corrosion coating to resist acidic water vapor corrosion, and there is a temperature sensor inside. It will automatically cut off the power when the temperature exceeds the limit to prevent fire.

[0067] Operating steps: Step 1: The pushing mechanism includes a support frame 2, a pushing motor 3, a splicing sleeve 4, an automatic lifting device 5, and a pushing guide plate 6. The support base 1 is used to support and prevent the equipment from bending and deforming. The upper part is connected to the support frame 2. The upper part of the support frame has a pushing motor 3, which is used to drive the automatic lifting device to work. The upper part of the automatic lifting device has a guide plate 6, which rises and falls with the automatic lifting device 5. Next, the upper part is the splicing sleeve 4. After it is fixed at the lower part, the lower device is spliced ​​with the splicing sleeve 4. Then, the explosives are put into the device, and the explosives are sent into the hole by the guide plate driven by the automatic lifting device.

[0068] Step 2: Both the splicing sleeve and the guide frame have guide grooves.

[0069] Step 3: The spraying mechanism includes a conveying pipe 7 in the splicing sleeve, a spraying device 8 at the end of the pipe, an interface 9 at the bottom of the pipe, a control valve 10 at the interface, and a replaceable filling device at the connection of the control valve. After the explosive is delivered to the designated position, the spraying mechanism starts to work. The replaceable filling device is spliced ​​with the interface 9, and the pipe connected inside the interface is tightly connected to the filling device. Then, driven by the motor 3, the device is pressurized, so that the filling material can smoothly reach the spraying device 8 through the pipe for spraying and filling. During the pushing process, driven by motor 3, automatic lifting device 5 and guide plate 6 slowly rise. When the medicine column is pushed to the designated position, the spraying mechanism starts spraying. As the lower pipe combines with the replaceable filling material, the replacement filling material is pressurized by motor 3 to smoothly reach the designated position through the pipe, completing the spraying. During the spraying process, the spraying speed can be adjusted by adjusting valve 10. As the medicine column slowly rises, the spraying device continues. When the medicine column is filled, automatic lifting device 5 and guide plate 6 slowly descend, and the spraying device continues. The bottom of the medicine column is then reinforced until the medicine column is completely integrated with the surrounding environment. After that, the filling is completed, power 11 is turned off, all devices are removed, and the process ends.

[0070] Step 4: Deployment of microseismic sensors and downhole monitoring.

[0071] Microseismic sensor placement: At least six microseismic sensors should be placed within a 1500m radius of the working face. Two sensors within the same tunnel should be 400-450m apart. Sensors closer to the working face should be moved if they are less than 100m away. The positions of the microseismic pickups should be adjusted promptly according to the actual site conditions to meet microseismic monitoring requirements. Figure 10 .

[0072] Microseismic event monitoring: Using a microseismic monitoring system, real-time monitoring of measuring points is carried out during the mining operation. The seismic source is located and the microseismic energy is calculated daily. The temporal-spatial-intensity evolution of microseismic activity is studied, and the impact hazard level of the working face area is evaluated.

[0073] Step 5: Evaluate the effect of the blasting to break the roof after the blast.

[0074] On-site blasting operations: (1) Determine the blasting construction parameters according to steps 1 to 4, and construct blasting holes at the designated locations according to the calculation results; (2) In order to achieve the expected blasting effect, this invention proposes to use the "energy-controlled blasting method" for on-site blasting.

[0075] Traditional blasting techniques often employ continuous charge structures. After detonation, the explosives in the blast hole detonate instantaneously, releasing a large amount of energy that acts on the rock strata instantly. The stress wave superposition effect is significant, which may cause excessive fragmentation of non-target rock strata, or even trigger mine tremors or shock disasters.

[0076] This invention relies on the aforementioned pre-splitting explosive charging device for the roof and adopts a construction approach of "spatial positioning based on the main fault → axial zoning → segmented filling → construction of discrete energy release units" to achieve an axially combined structure of "key reinforcement section + energy buffer section". The detailed construction steps include: ①Blast hole positioning: Perform spatial positioning and construction of blast holes according to steps 1-4; ② Axial Zoning Design of the Borehole: Along the axial direction of the borehole, from the borehole opening to the bottom (deepest point), the borehole is divided into: sealing section, ordinary charge section, energy control interval layer section, and key charge section. The key charge section should correspond to the main fault line to achieve targeted blasting. The zoning design is as follows: Figure 11 .

[0077] ③Segmented filling construction: a. Hole cleaning: Clean the rock debris inside the blast hole while ensuring the hole diameter remains intact; b. Explosive loading into the borehole: Explosives (including detonation components) are loaded into the roof pre-splitting blasting charge filling device. Explosives used in key charge sections are placed on top, while those used in ordinary charge sections are placed below. The main blasting target of key charge sections is the thick, hard rock layer of the main critical layer, using high explosives to generate stronger tensile stress. Ordinary charge sections are closer to the roadway and use medium- and low-temperature plastic-coated explosives to avoid fracturing the surrounding rock of the roadway. c. Installation of explosives in the key charge section: After the top plate pre-splitting blasting charge filling device transports the explosives to the designated location, the spraying device extends and fixes the explosives by grouting; two minutes later, the explosives in the key charge section are stabilized, and the top plate pre-splitting blasting charge filling device is retracted to the bottom of the key charge section. d. Construction of energy-controlled interval layers: After the pre-splitting blasting charge filling device in the top plate is withdrawn to the bottom of the key charge section, the spraying device sprays the filling material. The thickness of the interval layer is generally 0.3 to 1.0 times the length of the key charge section. After spraying, let it stand for two minutes. e. Installation of explosives in the ordinary charge section: After the construction of the energy-controlled interval layer is completed, the pre-splitting blasting charge filling device on the top plate is used to install the explosives used in the ordinary charge section after the interval layer, and the end spraying device is used to grout and fill and fix the explosives used in the ordinary charge section. f. Time-Graded Initiation: After confirming that the blasting environment meets safety regulations, the initiation components of the key charge section and the ordinary charge section are connected to the downhole initiation control system. The key charge section is set as the priority initiation unit, and the ordinary charge section is set as the delayed initiation unit. Finally, the initiation command is issued through the initiation control device, and each initiation unit detonates sequentially according to the set order, releasing the explosive energy in stages.

[0078] The present invention uses a weighted standardized comprehensive scoring method to evaluate the effect of blasting roof breaking. In the past, the evaluation of the effect of blasting roof breaking used the energy characteristics of microseismic events as a single evaluation index. The present invention, on the basis of the energy characteristics of microseismic events, considers the stability characteristics of the roof after blasting and innovatively introduces the roof period to press the step distance and the peak working resistance of the working face support as additional indicators, thus forming a multi-index evaluation system for the effect of blasting roof breaking.

[0079] The four factors for evaluating the effectiveness of blasting to break the roof are: a) the peak energy of micro-vibration events after blasting to break the roof; b) the number of micro-vibration events after blasting to break the roof; c) the peak working resistance of the working face support after blasting to break the roof; and d) the periodic pressure step distance after blasting to break the roof.

[0080] After automatic monitoring and analysis, the microseismic monitoring system obtains the energy peak value 'a' of the microseismic events following the blasting roof break and the number of microseismic events 'b'.

[0081] The hydraulic support pressure sensor at the working face records the phenomenon of sudden pressure increase during periodic pressure. Therefore, the peak working resistance c of the working face support after blasting and the periodic pressure step d after blasting and roof breakage can be obtained from this data.

[0082] Table 1. Classification of Indicator Meanings and Judgment of Target Trends

[0083] (2) Normalize the evaluation data to eliminate the influence of dimensions (J, times, kN, m).

[0084] Normalization is performed: ; ; ; .

[0085] In the formula, This represents the maximum energy value of the microseismic event. This represents the average energy of the microseismic events. This represents the minimum energy value of the microseismic event. This represents the maximum number of microseismic events. This represents the average number of microseismic events. This represents the minimum number of microseismic events. This represents the maximum working resistance of the support. This represents the average working resistance of the support. This represents the minimum working resistance of the support. The maximum step size is determined by the period; The average step size is used to control the periodicity. The minimum step size is determined by the periodicity. As the peak energy weight of microseismic events, As the weight of the number of microseismic events, As the peak resistance weight of the stent, The peak resistance of the stent is weighted.

[0086] The standardized main table values ​​are as follows:

[0087] Weighted composite score is used to evaluate the effectiveness of the top-breaking process: Evaluation score before blasting the roof: S 前 = ×ω a1 + ×ω b1 + ×ω c1 + ×ω d1 ; Evaluation score after blasting and breaking the roof: S 前 = ×ω a2 + ×ω b2 + ×ω c2 + ×ω d2 ; In the formula, This represents the average energy of micro-seismic events prior to the blasting and roof collapse. The average value of microseismic energy after blasting and roof breaking. This represents the average number of microseismic events prior to the blasting roof break. This represents the average number of microseismic events following the blasting and roof break. This represents the average working resistance of the support before the blasting roof break; This represents the average working resistance of the support after the blasting and roof break. The average step distance for periodic pressure application before blasting to break the roof; This represents the average step distance for periodic pressure application after blasting to break the roof. The weight of the peak energy of microseismic events before the blasting fracture is given. Weighting of the number of microseismic events prior to the blasting failure. The peak resistance of the support before the blasting roof break is weighted. The peak resistance of the support before the blasting roof break is weighted. The weight of the peak microseismic energy after blasting and roof breaking. Weighting of the number of microseismic events following blasting and roof collapse. The peak resistance of the support structure after blasting and roof breakage is weighted. The peak resistance of the support after the blasting and roof break is weighted.

[0088] Judgment formula: S = (S 后 / S 前 ) × 100%; S represents the proportion of residual energy in the roof after blasting. The smaller the S value, the better the blasting effect. When S≤60%, the blasting effect is excellent; when 60%≤S≤80%, the blasting effect is good; and when S≥80%, the blasting effect is poor.

[0089]

[0090] Example 1

[0091] Step 1: Collect geological data for a coal mine, including mine strata drilling data, whether there is a risk of rockburst in the surrounding mining area, and whether large-energy (10) events have occurred near the working face. 5J) Impact events and the risk of rockbursts in adjacent working faces. The report determines whether there is a risk of impact. Based on mine drilling data, a columnar section is created to divide the mine strata and determine the rock properties.

[0092] Step 2: Divide the main fault layer and the secondary collapse layer, and calculate the thickness of the rock strata in the theoretical collapse zone.

[0093] 1) The first main fault layer is determined. The load of the first rock layer itself is Q1 = 48.73 kPa. Considering the effect of the second layer on the first layer (W2)1, it is: (W2)1 = 60.23 kPa.

[0094] Considering the effect of layer 3 on layer 1 (W3)1, it is: (W3)1=38.69kPa, (W3)1<(W2)1.

[0095] Therefore, the first layer of siltstone serves as the main fault between layers 1 and 2, while the second layer of sandstone is a secondary caving layer. The total thickness is 2.45m.

[0096] 2) The second main fault layer is determined. The load of the third rock layer itself is Q3 = 65.17 kPa; considering the effect of the fourth layer on the third layer (W4)3, it is: (W4)3 = 81.80 kPa.

[0097] Considering the effect of layer 5 on layer 3 (W5)3, it is: (W5)3=35.72kPa, (W5)3<(W4)3.

[0098] Therefore, the third layer of siltstone serves as the main fault between layers 3 and 4, while layer 4 is a secondary caving layer. The total thickness is 3.25m.

[0099] 3) The third main fault layer is determined. The load on the fifth rock layer itself is Q5 = 101.29 kPa; considering the effect of the sixth layer on the fifth layer (W6)5, it is: (W6)5 = 138.31 kPa.

[0100] Considering the effect of layer 7 on layer 5 (W7)5, it is: (W7)5 = 162.79 kPa.

[0101] Considering the effect of layer 8 on layer 5 (W8)5, it is: (W8)5 = 197.19 kPa.

[0102] Consider the effect of layer 9 on layer 5 (W8)5, which is...

[0103] (W9)5=103.46kPa, (W9)5<(W8)5.

[0104] Therefore, the 5th layer of siltstone is the main fault layer for layers 5-8, and layers 6-8 are secondary collapse layers. The total thickness is 10.48m.

[0105] 4) The fourth main fault layer was determined. The load on the 9th rock layer itself is Q9 = 137.08 kPa; considering the effect of the 10th layer on the 9th layer (W) 10 )9, is: (W 10 )9=176.66kPa.

[0106] Considering the effect of layer 11 on layer 9 (W11)9, it is: (W 11 )9=89.38kPa, (W 11 )9<(W 10 9.

[0107] Therefore, the 9th layer of medium-grained sandstone is the main fault layer of layers 9 and 10, and layer 10 is the secondary collapse layer, with a total thickness of 6.95m.

[0108] The theoretical thickness of the collapsed rock layer is calculated using the formula: .

[0109] Since the upper strata of the mine roof are mainly hard sandstone, D = 5.42m and k are given in the formula. d =1.3~1.5.

[0110] The theoretical thickness of the collapsed rock layer was calculated to be 10.84~18.07 mm.

[0111] Step 3: L n Let L1 be the total thickness of the nth main fault layer and its corresponding secondary collapse layer. The relationship between the main fault layer, the secondary collapse layer, and the theoretical collapse layer thickness in the upper part of the mine roof is determined. The total thickness of the main fault layer 1 and its corresponding secondary collapse layer 1, L1 < Σh, so the determination continues upwards. L1 + L2 < Σh, so the determination continues upwards. L1 + L2 + L3 ≥ Σh, so the determination ends. Based on current deep-hole blasting technology in underground mines, the blasting drilling parameters are set, and the borehole is drilled to the 3rd main fault layer, with a blasting height h = 16.18m.

[0112] Explosive parameter settings: (1) Calculate the borehole angle according to the following formula. : ; .

[0113] The calculation yields: .

[0114] Pick , .

[0115] (2) Calculate the borehole depth based on geometric relationships : ; .

[0116] The calculation yields: .

[0117] .

[0118] Pick , .

[0119] (3) Calculate the charge amount per hole according to the following formula. : .

[0120] The blasting used emulsion explosives with a density ρ = 1300 kg / m³. 3 The length of the charge section is L1 = 23 / 3 ≈ 7.6m, L2 = 18 / 3 = 6m; ​​the diameter of the charge roll is d = 63mm.

[0121] The calculation yields: ; .

[0122] Pick , .

[0123] Hole arrangement: Holes are arranged in groups, with each group containing 4 holes in the goaf-side roadway and 3 holes in each group in the solid coal-side roadway. The borehole diameter is 80mm, and the group spacing is 15m. Roof blasting holes are constructed starting 40m away from the cut-off points of the two roadways and continuing to a position 50m away from the stop line.

[0124] Step 4: Several micro-vibration sensors are arranged in the roadways on both sides of the working face, and several micro-vibration sensors are installed on the mine hydraulic support. Each micro-vibration sensor receives the micro-vibration waveforms generated by the micro-vibration events generated at the working face and records them through the micro-vibration monitoring system.

[0125] Step 5: Evaluate the effect of the blasting to break the roof after the blast.

[0126] The blasting effect of the roof breakage in this area was evaluated based on the results recorded by the mine microseismic detection system.

[0127] (1) The monitoring results of the microseismic monitoring system are shown in the table below:

[0128] (2) Normalize a, b, c, and d: Normalization formula: ; ; ;

[0129] The calculation yields: , .

[0130] , .

[0131] , .

[0132] , .

[0133] (3) The standardized master table value table is as follows

[0134] (4) Weighted comprehensive score to evaluate the effect of the top break.

[0135] Pre-climb evaluation score: S 前 = ×ω a1 + ×ω b1 + ×ω c1 + ×ω d1 =1168.0109; Evaluation score after top breakage: S 前 = ×ω a2 + ×ω b2 + ×ω c2 + ×ω d2 =604.4144; Judgment formula: S = (S 后 / S 前 ) × 100% = 51.7%.

[0136] (5) Take appropriate measures based on the result S of the judgment formula.

[0137]

[0138] Data from the microseismic monitoring system shows that after the pre-cracking blasting of the roof, the peak value of microseismic energy decreased, the number of microseismic events decreased, the peak value of support resistance decreased, and the periodic pressure step distance decreased, indicating that the pre-cracking effect of the roof was excellent.

[0139] Of course, the above description is not intended to limit the present invention, and the present invention is not limited to the examples given above. Any changes, modifications, additions or substitutions made by those skilled in the art within the scope of the present invention should also fall within the protection scope of the present invention.

Claims

1. A roof blasting pressure relief method based on stratigraphic division and weighted comprehensive scoring, characterized in that, Includes the following steps: Step 1: Conduct on-site investigation to obtain geological condition parameters of the working face to be tested; Step 2: Divide the main fault and the secondary collapse zone, and calculate the theoretical thickness of the strata in the collapse zone; Step 3: Determine the blasting parameters, including determining the blasting height of the roof break, the arrangement of the blast holes, the blasting angle, the blast hole depth, and the charge per hole; Step 4: Deployment of microseismic sensors and downhole monitoring; Step 5: Evaluate the effect of the blasting to break the roof after the blast.

2. The roof blasting pressure relief method based on stratigraphic division and weighted comprehensive scoring according to claim 1, characterized in that, Step 2 specifically includes: There are several rock strata above the coal seam, which are, from bottom to top: the nth stratum, the (n+1)th stratum, ... the (n+i-1)th stratum, the (n+i)th stratum, and the (n+i+1)th stratum; Main fault layer division: The load applied by the (n+i+1)th rock layer to the nth rock layer ( W n+i+1 ) n The load exerted on the nth rock layer by the (n+i)th rock layer is less than that exerted on the nth rock layer. W n+i ) n Then the nth layer is the main fault layer; Sub-collapse layer classification: Layers from n+1 to n+i are considered sub-collapse layers; The load is calculated as follows: ； In the formula, W n The load of the nth layer itself is kN; The load of the nth rock layer on the 1st rock layer is expressed in kN. E The elastic modulus is expressed in MPa. The thickness of the nth rock layer; Let n be the unit weight of the nth rock stratum; The formula for calculating the theoretical caving zone thickness is: ； In the formula, D The effective mining height is in meters (m). Σh The theoretical thickness of the caving zone is in meters (m). This represents the average fracture expansion coefficient of the rock strata in the caving zone.

3. The roof blasting pressure relief method based on stratigraphic division and weighted comprehensive scoring according to claim 2, characterized in that, Step 3 specifically includes: The method for determining the blasting height of the fault is as follows: blast the main fault strata within the theoretical caving zone thickness range. That is, by determining the number and location of the main faults within the theoretical caving zone thickness, blasting is carried out at the height of the main fault locations. This method is divided into the following two types: (1) When the main fault layer n is determined, the height of the rock layer is equal to or exceeds the theoretical thickness of the collapsed rock layer. Σh In this case, the upper surface of the nth main fault layer is the highest point of the blasting height; (2) If the number and thickness of the main fault layers are too high and have not yet reached the theoretical thickness of the collapsed rock layers, then the fault is considered to be faulty. Σh The economic efficiency of blasting is too low. Taking into account the impact of main faults of different heights on mine pressure, the main faults with rock strata height within 30-50m above the coal seam are first blasted to break the top. Then, rockburst is prevented by a combination of measures, namely, pressure relief through coal seam drilling and strengthening of active and passive support of roadways. Arrangement of blast holes: The blast holes are arranged in groups. In solid coal roadways, there are 3 blast holes in each group; in goaf roadways, there are 4 blast holes in each group. The borehole diameter is 80+ mm and the group spacing is 15 m. The roof blasting holes are started 40 m away from the cut-out of the two roadways and continue to a position 50 m away from the stop line. Among them, the borehole angle of borehole No. 1 in the solid coal roadway The angle of borehole #2 The angle of borehole #3 It is 60°; Calculate according to the following formula : ； ； In the formula: The angle between the borehole axis and the horizontal plane; The actual height of the target main fault. The working face width is the length of the working face that advances along the dip of the coal seam. The distance between the tops of boreholes #1 and #2; The angle of the blast hole at blast hole #4 along the tunnel The borehole angle for borehole #5 is 40°~60°. The borehole angle for borehole #6 is 60°~70°. The borehole angle is 60° for borehole #7. It is 50°; The depth of each borehole was calculated based on geometric relationships. : ； m takes the integer value from 1 to 7; Calculate the charge amount per orifice using the following formula. : ； In the formula: Explosive density; For the length of the charge section, take... ; This refers to the cross-sectional area of ​​the explosive.

4. The roof blasting pressure relief method based on stratigraphic division and weighted comprehensive scoring according to claim 1, characterized in that, Microseismic sensor placement: At least 6 microseismic sensors shall be placed within a 1500m radius around the working face. Two microseismic sensors in the same tunnel shall be 400-450m apart. Microseismic sensors closer to the working face shall be moved if they are less than 100m away from the working face. The positions of the microseismic sensors shall be adjusted in a timely manner according to the actual site conditions to meet the microseismic monitoring requirements. Microseismic event monitoring: Using a microseismic monitoring system, the measuring points are monitored in real time during the mining operation, and the seismic source is located and the microseismic energy is calculated every day.

5. The roof blasting pressure relief method based on stratigraphic division and weighted comprehensive scoring according to claim 1, characterized in that, Step 5 specifically includes: The peak energy value 'a' of the microseismic events after the blasting roof break was obtained, and the number of microseismic events 'b' of the blasting roof break was obtained. The peak working resistance c of the working face support after the blasting roof break is obtained, and the periodic pressure step d after the blasting roof break is obtained; Normalization is performed: ； ； ； ； In the formula, This represents the maximum energy value of the microseismic event. This represents the average energy of the microseismic events. This represents the minimum energy value of the microseismic event. This represents the maximum number of microseismic events. This represents the average number of microseismic events. This represents the minimum number of microseismic events. This represents the maximum working resistance of the support. This represents the average working resistance of the support. This represents the minimum working resistance of the support. The maximum step size is determined by the period. The average step size is used to control the periodicity. The minimum step size is determined by the periodicity. As the peak energy weight of microseismic events, As the weight of the number of microseismic events, As the peak resistance weight of the stent, The peak resistance of the stent is weighted. Evaluation score before blasting the roof: S 前 = ×ω a1 + ×ω b1 + ×ω c1 + ×ω d1 ; Evaluation score after blasting and breaking the roof: S 前 = ×ω a2 + ×ω b2 + ×ω c2 + ×ω d2 ; In the formula, This represents the average energy of micro-seismic events prior to the blasting and roof collapse. The average value of microseismic energy after blasting and roof breaking. This represents the average number of microseismic events prior to the blasting roof break. This represents the average number of microseismic events following the blasting and roof break. This represents the average working resistance of the support before the blasting roof break; This represents the average working resistance of the support after the blasting and roof break. The average step distance for periodic pressure application before blasting to break the roof; This represents the average step distance for periodic pressure application after blasting to break the roof. The weighting of the peak energy of microseismic events before the blasting fracture is given. Weighting of the number of microseismic events prior to the blasting failure. The peak resistance of the support before the blasting roof break is weighted. The peak resistance of the support before the blasting roof break is weighted. The weight of the peak microseismic energy after blasting and roof breaking. Weighting of the number of microseismic events following blasting and roof collapse. The peak resistance of the support structure after blasting and roof breakage is weighted. The peak resistance of the support after blasting and roof breakage is weighted. Judgment formula: S = (S 后 / S 前 ) × 100%; S represents the proportion of residual energy in the roof after blasting. The smaller the S value, the better the blasting effect. When S≤60%, the blasting effect is excellent; when 60%≤S≤80%, the blasting effect is good; and when S≥80%, the blasting effect is poor.

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