A mining stress regulation method during a shallow-buried fully-mechanized mining face final mining period

Abstract

This invention discloses a method for controlling mining-induced stress during the final stage of shallow-buried fully mechanized mining faces, relating to the field of coal mining technology. The method includes: acquiring overlying geological parameters and calculating the theoretical fracture step distance of the basic roof; establishing a numerical model to determine the instability mode and the target fracture location of forced roof caving; when the working face advances to the critical range and meets multiple indicator judgment conditions, entering the stress control window period and stopping mining to reduce pressure; after pressure is applied, forced roof caving is implemented at the target fracture location; based on the decrease in stress concentration coefficient and the reduction in the high-stress zone area after forced roof caving, the support parameters are dynamically adjusted in stages and verified; after successful verification, the advance is resumed with a graded reduction in speed until breakthrough is achieved. This invention achieves precise and proactive control of the timing and location of fracture through the coordinated control of stopping mining to reduce pressure and forced roof caving, and establishes a closed-loop dynamic control mechanism, effectively solving problems such as large deformation of the surrounding rock, high pressure, and difficult support in the retreat channel during the final stage of mining.

Description

Technical Field

[0001] This invention relates to the field of coal mining technology, specifically a method for controlling mining-induced stress during the final stage of mining in a shallow fully mechanized longwall face. Background Technology

[0002] The coal seams in western my country's mining areas are characterized by shallow burial, large thickness, small dip angle, and thin bedrock. To accelerate the continuity of the working face, pre-excavated double retreat channels are widely used in engineering practice, combined with trackless rubber-tired vehicles for transportation. When the working face advances to near the main retreat channel, the surrounding rock of the main retreat channel undergoes severe deformation due to the combined effects of mining-induced stress and lateral support pressure. This often results in significant roof subsidence and severe spalling of the solid coal face, seriously threatening safe production during the final stages of mining.

[0003] To address the aforementioned problems, existing solutions mainly include full-section filling reinforcement, single-blast pre-splitting, and mining height and advance speed adjustment. While full-section filling reinforcement can control surrounding rock deformation to some extent, it involves massive engineering scale and a long construction period. Single-blast pre-splitting relies solely on this method, lacking proactive control over the timing of the basic roof breakage. When the timing of blast pre-splitting and the pressure from the working face do not match, strong pressure may still occur during the breakthrough stage. The mining height and advance speed adjustment method faces the dual contradiction of production efficiency and subsequent pressure. Existing technologies still have technical shortcomings in solving the problem of surrounding rock control in the final stage of shallow-buried fully mechanized mining faces with double-retreat channels: First, they lack the ability to coordinate the timing and location of basic roof breakage; second, a systematic technical process for mining-induced stress control has not been established; third, a dynamic control mechanism based on field measurements is lacking; and fourth, support design and stress control are disconnected. Therefore, there is an urgent need to propose a synergistic mining-induced stress system control method. Summary of the Invention

[0004] The purpose of this invention is to provide a method for controlling mining-induced stress during the final mining phase of a shallow-buried fully mechanized mining face, thereby overcoming the aforementioned problems existing in the prior art.

[0005] To achieve the objective, the present invention provides the following technical solution: A method for controlling mining-induced stress during the final stage of mining in a shallow-buried fully mechanized longwall face includes the following steps: Step 1: Obtain the geological parameters of the overlying strata on the working surface. The geological parameters include at least the thickness, uniaxial tensile strength, and elastic modulus of the basic top stratum, as well as the cumulative thickness and average density of the strata above the basic top. Step 2: Based on the geological parameters, establish a cantilever deep beam mechanical model and calculate the theoretical fracture step distance of the basic roof; establish a stope-retreat channel numerical model to simulate the fracture process of the basic roof in the final mining stage, and determine the instability form of the basic roof based on the relative position of the fracture line of the basic roof and the main retreat channel, and determine the key control distance and the target fracture position of forced roof caving. Step 3: When the working face advances to the first preset distance before the theoretical fracture step distance, the monitoring system is activated; when the working face advances to the range of the second preset distance before the theoretical fracture step distance to the theoretical fracture step distance, and meets the preset judgment conditions, it is determined that the stress control window period has been entered; wherein, the first preset distance is greater than the second preset distance; Step 4: After entering the stress control window period, stop the working face advance, maintain the rated working resistance of the support, monitor the mine pressure manifestation parameters, and wait for the basic roof to complete one full pressure injection. Step 5: After the pressure is applied, a forced jacking is performed at the target fracture location to cause the basic top edge to fracture at the predetermined cross-section. Step 6: After forced roof collapse, monitor the stress distribution changes within the coal pillar, calculate the stress transfer effect index, and adjust the support parameters of the main retreat channel in stages according to the stress transfer effect index until the support effect is verified as qualified. Step 7: After the support effect is verified to be qualified, resume the advancement of the working face until the working face is connected to the main withdrawal channel.

[0006] Furthermore, in step two, the formula for calculating the theoretical fracture step distance is: in, This represents the basic thickness of the overlying strata. The tensile strength of the overlying rock mass is the basic tensile strength of the rock mass. For load transfer factor, The average density of the overlying strata is the basic overlying strata. The cumulative thickness of the overlying strata is the basic top layer.

[0007] Furthermore, in step three, the first preset distance is 30m, the second preset distance is 5m, and the judgment conditions include: the working resistance of the hydraulic support of the working face continues to rise, and the rise exceeds 20% of the normal value; the roof sinking rate of the main retreat channel exceeds 10mm / d for 2 consecutive hours; and the stress concentration coefficient of the coal pillar between the main retreat channels reaches 1.5 or more.

[0008] Furthermore, in step four, the conditions for determining the end of the pressure run include at least: the working resistance of the support decreases to within 1.2 times the normal value, and the rate of top plate sinking is less than 5 mm / d for 4 consecutive hours.

[0009] Furthermore, in step five, the method for determining the target fracture location includes: When the instability mode predicted by numerical simulation is sliding instability, the optimal fracture location is determined by comparing multiple indicators through numerical simulation and calculation. The first preferred indicator is the minimum stress concentration coefficient of the coal pillar, the second preferred indicator is the minimum roof subsidence of the main retreat channel, and the third preferred indicator is the minimum peak working resistance of the support. When the instability mode predicted by numerical simulation is rotational instability, the optimal fracture location is determined by comparing multiple indicators, with the minimum rotation angle as the first preferred indicator, the minimum rotational angular velocity as the second preferred indicator, and the minimum horizontal displacement of the top plate of the pullback channel as the third preferred indicator.

[0010] Furthermore, in step five, the forced caving adopts a directional hydraulic fracturing method, in which fracturing boreholes are drilled in the basic roof rock stratum, the boreholes penetrate the full thickness of the basic roof, and a continuous pre-fracturing surface is formed at a predetermined cross-section by constant pressure water injection.

[0011] Furthermore, in step six, the stress transfer effect indicators include the decrease in the stress concentration factor and the reduction in the range of the high-stress area. When the decrease in stress concentration factor is greater than 30% and the reduction in the range of high stress area is greater than 40%, a conventional support scheme is adopted; when the decrease in stress concentration factor is 20% to 30% and the reduction in the range of high stress area is 30% to 40%, a reinforced support scheme is adopted; when the decrease in stress concentration factor is less than 20% or the reduction in the range of high stress area is less than 30%, a strengthened support scheme is adopted.

[0012] Furthermore, in step seven, a graded speed reduction method is adopted when resuming advance: after resuming advance, the coal mining machine advances at 50% of the normal speed, pausing to observe every 2m; when it advances to 5m from the main retreat channel, the speed is reduced to 30% of the normal speed, pausing every 1m; when it advances to 2m from the main retreat channel, the coal mining machine is stopped, and the breakthrough is completed manually.

[0013] Compared with the prior art, the present invention has the following advantages: Firstly, this invention employs a holistic approach, from steps one to seven, to acquire overburden parameters, calculate theoretical fracture distances, and establish a numerical model to determine instability patterns. Then, based on multi-index monitoring, it dynamically identifies stress control windows. Within these windows, it sequentially executes measures such as stopping mining to reduce pressure, forced roof caving, stress transfer followed by tiered dynamic support, and tiered deceleration for breakthrough. This achieves coordinated and proactive control over the timing and location of basic roof fracture, transforming passive response into proactive pre-control. It fundamentally avoids severe mine pressure manifestations caused by slippage or rotational instability during breakthrough. Simultaneously, it establishes a closed-loop dynamic control mechanism encompassing monitoring, pressure reduction, roof caving, and support, significantly improving the reliability and intelligence of surrounding rock control in the final mining stage.

[0014] Secondly, the theoretical fracture step distance calculation formula in step two of this invention is based on the cantilever deep beam mechanical model and the maximum tensile stress criterion, quantifying key parameters such as the basic top thickness, tensile strength, and overlying strata load. This formula provides a precise quantitative basis for determining the theoretical fracture step distance, significantly improving calculation accuracy compared to the empirical analogy method, and laying a scientific mechanical foundation for determining the subsequent monitoring window period and yield point.

[0015] Thirdly, in step three of this invention, the starting distance of the monitoring system is set as the first preset distance before the theoretical fracture step distance, and the determination conditions for the stress control window period are specifically quantified as follows: the working resistance of the support increases by more than 20%, the roof subsidence rate exceeds 10 mm / d for two consecutive hours, and the stress concentration coefficient of the coal pillar reaches 1.5 or higher. This multi-indicator fusion determination mechanism effectively avoids the drawbacks of misjudgment or lag of a single indicator, and can accurately capture the key time window before the basic roof is about to be crushed, ensuring that the control measures are initiated at the most favorable time, and can significantly improve the timeliness and effectiveness of the control.

[0016] Fourth, in step four of this invention, the criterion for determining the end of the pressure release is clearly defined as the working resistance of the support decreasing to within 1.2 times the normal value and the rate of roof sinking being less than 5 mm / d for 4 consecutive hours. This quantitative standard is based on actual on-site measurement data, which can accurately control the pressure release termination time, ensuring that the elastic energy accumulated in the basic roof is fully released without excessively prolonging the downtime, thus achieving a good balance between safety and efficiency.

[0017] Fifth, in step five of this invention, differentiated multi-index optimization methods are used to determine the target fracture location for two different expected instability modes: sliding instability and rotational instability. For sliding instability, the minimum stress concentration coefficient of the coal pillar is the primary index; for rotational instability, the minimum rotation angle is the primary index. This optimization strategy tailored to the instability mode ensures that the fracture line after forced roof caving is precisely located in the position most conducive to the stability of the surrounding rock, minimizing the pressure intensity or rotational impact during breakthrough.

[0018] Sixth, in step six of this invention, the support scheme is divided into three levels—conventional support, reinforced support, and enhanced support—based on two quantitative indicators: the decrease in stress concentration coefficient and the reduction in the range of high-stress areas. This graded response mechanism enables the support parameters to be precisely matched with the actual stress transfer effect after forced caving, achieving dynamic synergistic optimization of the support system and effectively solving the problem of disconnect between support design and stress state in existing technologies. Furthermore, in step seven of this invention, a graded deceleration method is adopted after resuming advance. This gradual advance strategy ensures a smooth transition of surrounding rock stress, avoiding additional dynamic load disturbances caused by sudden speed changes. Attached Figure Description

[0019] Figure 1 This is a flowchart of the mining-induced stress control method for the final mining stage of a shallow-buried fully mechanized mining face according to the present invention.

[0020] Figure 2 This is a mechanical model diagram of a cantilever deep beam subjected to a uniformly distributed load in this invention. In the diagram, This represents the basic thickness of the overlying strata. For load transfer factor, The average density of the overlying strata is the basic overlying strata. The cumulative thickness of the overlying strata is the basic top layer.

[0021] Figure 3 This is a schematic diagram of the pressure relief position in this invention. In the diagram, D1 is the width of the pressure relief interval coal pillar, and S is the width of the main retraction channel.

[0022] Figure 4 This is a schematic diagram of the roof structure when the working face and the main retreat channel are connected in this invention. In the diagram, L is the periodic pressure step distance, D1 is the width of the pressure relief interval coal pillar, S is the width of the main retreat channel, and D2 is the width of the pressure relief interval coal pillar to ensure no pressure is applied during connection. Detailed Implementation

[0023] Specific embodiments of the present invention will now be described with reference to the accompanying drawings. Many details are described below to provide a comprehensive understanding of the invention; however, those skilled in the art will be able to implement the invention without these details.

[0024] like Figures 1 to 4 As shown, a method for controlling mining-induced stress during the final stage of mining in a shallow-buried fully mechanized longwall face includes the following specific steps: Step 1: Geological exploration and acquisition of overlying parameters.

[0025] Based on the mine geological exploration data and borehole columnar sections near the working face, the lithological combination, thickness distribution, and physical and mechanical parameters of the overlying strata on the working face are determined. The focus is on obtaining parameters such as the thickness, uniaxial tensile strength, elastic modulus, and Poisson's ratio of the basic roof strata, as well as the cumulative thickness and average density of all strata above the basic roof to the surface. Simultaneously, basic information such as coal seam thickness and mining height is recorded.

[0026] Step 2: Analysis of the basic top fracture characteristics and judgment of instability mode.

[0027] (1) Based on the geological parameters explored in Step 1, the basic top is regarded as a cantilever beam structure subjected to uniformly distributed load, and a model is established as follows: Figure 2 The mechanical analysis model shown is the cantilever deep beam mechanical model. Based on the stress distribution characteristics of beams in elasticity mechanics and combined with the maximum tensile stress criterion, the calculation formula for the ultimate span (i.e., theoretical failure step) of the basic top cantilever deep beam at failure is derived as follows: in, The thickness of the basic top of the overlying strata is expressed in meters (m). The tensile strength of the overlying rock mass is given in MPa. The load transfer factor is set to 0.5. The average density of the overlying strata is taken as 2500 kg / m³. 3 ; The cumulative thickness of the overlying strata is given in meters (m).

[0028] (2) A coupled numerical model of the stope-retreat channel was established using 3DEC numerical simulation software to simulate the entire process of the working face moving from away from the retreat channel to breakthrough. Data such as the stress distribution of the surrounding rock, the displacement of the roof, and the location of the basic roof failure were recorded.

[0029] Extracting the location of the basic roof fracture line at the moment of breakthrough from the numerical simulation results, and judging the expected instability form of the basic roof at the moment of breakthrough based on the relative positional relationship between the fracture line and the main retreat channel: if the fracture line is located directly above the main retreat channel, it is judged as slippage instability; if the fracture line is located above the coal pillar of the main retreat channel, it is judged as rotation instability.

[0030] (3) Based on the fracture location of the basic top recorded in the numerical simulation and the advance distance of the working face, the actual fracture step distance of the basic top is calculated. The actual fracture step distance is compared with the theoretical fracture step distance obtained from the mechanical model. The smaller value is taken as the lower limit of the theoretical fracture step distance and the larger value is taken as the upper limit of the theoretical fracture step distance, thereby determining the numerical range of the theoretical fracture step distance.

[0031] Extract the vertical stress distribution data within the coal pillar at each simulated step distance, and calculate the ratio of the maximum vertical stress within the coal pillar in the main retreat channel to the original rock stress, using this as the stress concentration factor. Regions with a stress concentration factor greater than 1.2 are marked as high-stress zones, and the location coordinates and size of these high-stress zones are recorded at different advance distances. The critical advance distance at which the stress concentration factor begins to rise significantly is determined as the key control distance.

[0032] Based on the aforementioned anticipated instability patterns, the target break point for forced top release is set as follows: (a) When the expected instability is slippage instability, the candidate area for the target fracture location is set above the coal pillar of the main retreat channel. In the candidate area, a series of preset fracture lines at different distances from the edge of the main retreat channel are set along the working face direction using numerical simulation software. The three quantitative indicators of coal pillar stress concentration coefficient, main retreat channel roof subsidence, and support working resistance peak value are simulated when each preset fracture line along the basic roof fractures. The minimum coal pillar stress concentration coefficient is used as the first preferred indicator, the minimum roof subsidence is used as the second preferred indicator, and the minimum support working resistance peak value is used as the third preferred indicator. The optimal fracture location is determined by comprehensive comparison of multiple indicators and is used as the final target fracture location.

[0033] As a preferred option, the stress concentration coefficient of the coal pillar corresponding to each preset break line is compared, and the preset break line with the smallest stress concentration coefficient is selected as the candidate location. When there are multiple preset break lines with the same stress concentration coefficient or a difference of less than 5% at the candidate location, the roof subsidence of the main retreat channel corresponding to these preset break lines is compared, and the preset break line with the smallest roof subsidence is selected as the candidate location. When there are still preset break lines with the same roof subsidence or a difference of less than 5% at the candidate location, the preset break line with the smallest peak working resistance of the support is selected as the optimal break location.

[0034] (b) When the expected instability is rotational instability, the candidate area for the target fracture location is set as far away from the retreat channel as possible to reduce the impact of the rotational moment on the retreat channel. In the candidate area, a series of preset fracture lines at different distances from the edge of the main retreat channel are set along the working face using numerical simulation software. The rotational instability process of the basic top corresponding to each preset fracture line is simulated, and three quantitative indicators are extracted: rotational angle, rotational angular velocity, and horizontal displacement of the top plate of the retreat channel. The minimum rotational angle is used as the first preferred indicator, the minimum rotational angular velocity is used as the second preferred indicator, and the minimum horizontal displacement of the top plate is used as the third preferred indicator. The optimal fracture location is determined by comprehensive comparison of multiple indicators and is used as the final target fracture location.

[0035] Preferably, the rotation angles corresponding to each preset break line are compared, and the preset break line with the smallest angle is selected as the candidate position. If multiple preset break lines in the candidate position have the same rotation angle or a difference of less than 5%, the rotation angular velocities corresponding to these preset break lines are compared, and the preset break line with the smallest rotation angular velocity is selected as the candidate position. If multiple preset break lines in the candidate position still have the same rotation angular velocity or a difference of less than 5%, the horizontal displacement of the top plate of the retraction channel corresponding to these preset break lines is compared, and the preset break line with the smallest horizontal displacement of the top plate is selected as the optimal break position.

[0036] Step 3: Deployment of monitoring system and determination of stress control window period.

[0037] When the working face advances to the first preset distance before the theoretical fracture step, the monitoring system is activated. When the working face advances to the range between the second preset distance before the theoretical fracture step and the theoretical fracture step, and meets the preset judgment conditions, it is determined that the stress control window period has begun; if not, the working face continues to advance, and the monitoring frequency is increased. The first preset distance is greater than the second preset distance.

[0038] The monitoring system is specifically deployed as follows: A roof delamination meter is installed every 10m along the roadway axis on the roof of the main retreat passage to monitor roof subsidence and its rate; a set of displacement monitoring points is installed every 15m along the solid coal side of the main retreat passage, each set including surface displacement gauges and deep displacement gauges to monitor surface approach and internal deformation depth, respectively; a set of pressure sensors is pre-embedded every 20m within the coal pillar between the main and auxiliary retreat passages, each set including three pressure sensors, located at the edge, middle, and opposite edge of the coal pillar to monitor the distribution and variation of vertical stress within the coal pillar; pressure sensors are installed on the hydraulic supports of the working face to collect real-time support resistance data; microseismic monitoring probes are arranged in a grid pattern behind the goaf and around the retreat passage, with a grid spacing of 50m, to monitor the frequency and energy release characteristics of microseismic events. All monitoring equipment is integrated through an intrinsically safe mine data acquisition substation and transmitted in real-time to the ground dispatch center.

[0039] After the monitoring system is deployed, a stress control window period determination mechanism based on multi-source information fusion is established. Using the face advance step distance as the time coordinate, the following key monitoring indicators are recorded and analyzed in real time: the changing trend of the working resistance of the hydraulic supports at the working face, calculating the increase in working resistance relative to the normal advance stage; the subsidence rate of the roof in the main retreat channel, calculating the subsidence per unit time using continuous monitoring data from the roof separation instrument; and the stress concentration factor of the coal pillar, i.e., the ratio of the maximum vertical stress within the coal pillar to the original rock stress. Based on the critical advance distance in step two, combined with the basic roof theory fracture step distance, the distance range of the stress control window period is determined.

[0040] In one specific embodiment, the first preset distance is 30m, and the second preset distance is 5m. At this time, when the working face advances to ( When the working face advances to +30m, the monitoring system is activated; when the working face advances to +5m to When the stress is within the specified range and meets the preset judgment conditions, it is determined that the stress control window period has begun.

[0041] In one specific embodiment, the following three conditions must be met simultaneously to enter the stress control window period: (1) the working resistance of the support continues to rise and the rise exceeds 20% of the normal value; (2) the roof sinking rate exceeds 10 mm / d for 2 consecutive hours; (3) the stress concentration factor of the coal pillar reaches 1.5 or more.

[0042] During the judgment process, the monitoring data is input into the preset data processing system. The data processing system automatically generates the change curve of the monitoring indicators and dynamically compares it with the preset threshold. When all three indicators reach or exceed the threshold (i.e., the above three judgment conditions are met at the same time), the system automatically issues a control start command.

[0043] Step 4: Stop mining and reduce pressure.

[0044] After entering the stress control window, stop advancing the working face, maintain the rated working resistance of the support, monitor the mine pressure manifestation parameters, and wait for the basic roof to complete a full pressure cycle. The core purpose of the stop-mining and pressure-yielding operation is to allow the basic roof to complete a full pressure cycle within the existing fracture step distance, release the accumulated elastic energy in advance, and avoid strong pressure when the working face and the main retreat channel are connected.

[0045] The specific implementation method for the shutdown and pressure relief operation is as follows: After entering the stress control window period, the coal mining machine is shut down and the power is cut off. The support operator adjusts the working face support to an active support state, maintaining the rated working resistance of the support, and does not perform any propulsion-related operations. During the shutdown period, the monitoring system continues to operate normally, and the data acquisition frequency is increased from 10 minutes / time in the normal stage to 2 minutes / time. The following indicators are monitored in detail: the change curve of the support working resistance, the roof subsidence rate, the frequency and energy of micro-vibration events, and the vertical stress distribution in the coal pillar.

[0046] When the working resistance starts to rise continuously from the normal value, it indicates that the pressure has begun. A complete pressure is determined to be completed when the following preset judgment conditions are met simultaneously: (1) the working resistance of the support drops to within 1.2 times the normal value; (2) the roof subsidence rate is less than 5 mm / d for 4 consecutive hours; (3) the frequency of microseismic events drops to within 1.5 times the background level; (4) there is no energy greater than 1×10 for 6 consecutive hours. 4 J's incident occurred.

[0047] As a preferred option, an emergency response mechanism should be established during the pressure relief period. Emergency control should be initiated immediately when any of the following abnormal situations occur: (1) The working resistance of the support increases by more than 50% of the normal value and continues to rise without any downward trend; (2) The roof subsidence rate exceeds 20 mm / d for 2 consecutive hours and a local high-stress zone with a stress concentration factor exceeding 2.0 appears in the coal pillar; (3) Microseismic monitoring shows a single energy greater than 1×10 5 J is a high-energy event. When any of the above-mentioned abnormal situations occur, the on-site personnel should immediately evacuate to a safe area. At the same time, the ground dispatch center should activate the emergency plan, and carry out high-pressure grouting reinforcement of the coal pillar through the pre-set backup grouting boreholes, or implement pressure relief measures through the pre-buried pressure relief boreholes. After the monitoring data returns to normal, the decision on whether to continue to reduce pressure or adjust the control plan will be reassessed.

[0048] Step 5: Force the top to control the breaking position.

[0049] After the pressure is released, a forced caving operation is initiated, and the forced caving is carried out at the target fracture location to cause the basic roof edge to fracture at the predetermined cross section.

[0050] As a preferred method, based on the target fracture location determined in step two, a directional hydraulic fracturing method is employed to drill fracturing boreholes in the top rock stratum. The boreholes penetrate the full thickness of the top rock stratum, and a continuous pre-fracture surface is formed at a predetermined cross-section by constant-pressure water injection. The location of the fracture surface is verified by microseismic monitoring or borehole inspection. If the deviation exceeds the allowable range, additional fracturing is performed.

[0051] In one specific embodiment, fracturing boreholes are drilled in the overlying rock strata, penetrating the full thickness of the overlying rock and extending 1–2 m into the overlying strata, with a borehole spacing of 10–15 m. After installing the borehole sealer, water is injected at an initial pressure of 5–10 MPa, gradually increasing to the design fracturing pressure of 15–25 MPa. Constant water pressure is maintained until the rock strata fracture. The direction of fracture propagation is monitored by microseismic monitoring. Each borehole is fracturing sequentially, forming a continuous pre-fractured surface along the working face, causing the overlying rock to fracture along a predetermined cross-section.

[0052] Step Six: Coordination of Mining Stress Transfer and Support.

[0053] After forced roof collapse, monitor the stress distribution changes within the coal pillar, calculate the stress transfer effect index, and adjust the support parameters of the main retreat channel in stages according to the stress transfer effect index until the support effect is verified as qualified.

[0054] In one specific embodiment, after forced roof caving is completed, the change in vertical stress distribution within the coal pillar is continuously monitored. Using the average value of the hour prior to forced roof caving as a baseline, and the average value of the 24 hours following forced roof caving as the change value, the decrease in the stress concentration factor is calculated. and the reduction in the range of high-stress areas .

[0055] The decrease in stress concentration factor The calculation formula is:

[0056] in, The baseline stress concentration factor is the ratio of the maximum vertical stress in the coal pillar to the original rock stress before forced roof caving. The stress concentration factor is a variable value, which is the ratio of the maximum vertical stress in the coal pillar to the original rock stress after forced roof caving.

[0057] The extent of reduction in the range of high-stress areas The calculation formula is:

[0058] in, Before forced caving, the area (m) with a stress concentration factor greater than 1.2 is the high stress zone of the reference value. The range (m) of stress concentration factor greater than 1.2 after forced caving is defined as the range of high stress variation.

[0059] Adjust support parameters according to the calculation results in stages: when >30% and When >40%, a conventional support scheme is adopted; when 20% to 30% and When the area is 30%–40%, a reinforced support scheme should be adopted; when… <20% or When the coal seam thickness is less than 30%, a reinforced support scheme is adopted. This reinforced support scheme includes grouting reinforcement using grouting anchors in the roof construction, grouting of the side walls using grouting holes in the solid coal side roadway, and the addition of stacked supports in the main withdrawal passage.

[0060] After the support is adjusted, continuous monitoring is conducted for 48 hours. The support effect is deemed qualified when the roof subsidence rate is <2mm / d, the side wall approach rate is <1mm / d, the coal pillar stress concentration factor is <1.5, and the working resistance fluctuation of the support is <10%.

[0061] Step 7: Gradually reduce speed to complete the breakthrough operation.

[0062] After the support effect is verified to be satisfactory, the working face is resumed to advance until the working face is connected to the main withdrawal channel.

[0063] In one specific embodiment, a graded speed reduction method is adopted when resuming advance: after resuming advance, the coal mining machine first advances at 50% of the normal speed, and pauses to observe every 2m; when it advances to 5m away from the main retreat channel, the speed is reduced to 30% of the normal speed, and pauses every 1m; when it advances to 2m away from the main retreat channel, the coal mining machine is stopped, and the breakthrough is completed manually.

[0064] The above are merely specific embodiments of the present invention, but the design concept of the present invention is not limited thereto. Any non-substantial modifications made to the present invention using this concept shall be considered as infringing upon the protection scope of the present invention.

Claims

1. A mining stress regulation method during the end-mining period of a shallow-buried fully-mechanized mining face, characterized in that, Includes the following steps: Step 1: Obtain the geological parameters of the overlying strata on the working surface. The geological parameters include at least the thickness, uniaxial tensile strength, and elastic modulus of the basic top stratum, as well as the cumulative thickness and average density of the strata above the basic top. Step 2: Based on the geological parameters, establish a cantilever deep beam mechanical model and calculate the theoretical fracture step distance of the basic roof; establish a stope-retreat channel numerical model to simulate the fracture process of the basic roof in the final mining stage, and determine the instability form of the basic roof based on the relative position of the fracture line of the basic roof and the main retreat channel, and determine the key control distance and the target fracture position of forced roof caving. Step 3: When the working face advances to the first preset distance before the theoretical fracture step distance, the monitoring system is activated; when the working face advances to the range of the second preset distance before the theoretical fracture step distance to the theoretical fracture step distance, and meets the preset judgment conditions, it is determined that the stress control window period has been entered; wherein, the first preset distance is greater than the second preset distance; Step 4: After entering the stress control window period, stop the working face advance, maintain the rated working resistance of the support, monitor the mine pressure manifestation parameters, and wait for the basic roof to complete one full pressure injection. Step 5: After the pressure is applied, a forced jacking is performed at the target fracture location to cause the basic top edge to fracture at the predetermined cross-section. Step 6: After forced roof collapse, monitor the stress distribution changes within the coal pillar, calculate the stress transfer effect index, and adjust the support parameters of the main retreat channel in stages according to the stress transfer effect index until the support effect is verified as qualified. Step 7: After the support effect is verified to be qualified, resume the advancement of the working face until the working face is connected to the main withdrawal channel.

2. The method for controlling mining-induced stress during the final stage of shallow-buried fully mechanized mining as described in claim 1, characterized in that, In step two, the formula for calculating the theoretical fracture step distance is: in, This represents the basic thickness of the overlying strata. The tensile strength of the overlying rock mass is the basic tensile strength of the rock mass. For load transfer factor, The average density of the overlying strata is the basic overlying strata. The cumulative thickness of the overlying strata is the basic top layer.

3. The method for controlling mining-induced stress during the final stage of shallow-buried fully mechanized mining as described in claim 1, characterized in that, In step three, the first preset distance is 30m and the second preset distance is 5m. The judgment conditions include: the working resistance of the hydraulic support of the working face continues to rise and the rise exceeds 20% of the normal value; the roof sinking rate of the main retreat channel exceeds 10mm / d for 2 consecutive hours; and the stress concentration coefficient of the coal pillar between the main retreat channels reaches 1.5 or more.

4. The method for controlling mining-induced stress during the final stage of shallow-buried fully mechanized mining as described in claim 1, characterized in that, In step four, the conditions for determining the end of the pressure release include at least: the working resistance of the support decreases to within 1.2 times the normal value, and the rate of top plate sinking is less than 5 mm / d for 4 consecutive hours.

5. The method for controlling mining-induced stress during the final stage of shallow-buried fully mechanized mining as described in claim 1, characterized in that, In step five, the method for determining the target fracture location includes: When the instability mode predicted by numerical simulation is sliding instability, the optimal fracture location is determined by comparing multiple indicators through numerical simulation and calculation. The first preferred indicator is the minimum stress concentration coefficient of the coal pillar, the second preferred indicator is the minimum roof subsidence of the main retreat channel, and the third preferred indicator is the minimum peak working resistance of the support. When the instability mode predicted by numerical simulation is rotational instability, the optimal fracture location is determined by comparing multiple indicators, with the minimum rotation angle as the first preferred indicator, the minimum rotational angular velocity as the second preferred indicator, and the minimum horizontal displacement of the top plate of the pullback channel as the third preferred indicator.

6. The method for controlling mining-induced stress during the final stage of shallow-buried fully mechanized mining faces according to claim 1 or 5, characterized in that, In step five, the forced caving adopts the directional hydraulic fracturing method, in which fracturing boreholes are drilled in the basic roof rock layer, the boreholes penetrate the full thickness of the basic roof, and a continuous pre-fracturing surface is formed at the predetermined section by constant pressure water injection.

7. The method for controlling mining-induced stress during the final stage of shallow-buried fully mechanized mining as described in claim 1, characterized in that, In step six, the stress transfer effect indicators include the decrease in stress concentration factor and the reduction in the range of high stress area. When the decrease in stress concentration factor is greater than 30% and the reduction in the range of high stress area is greater than 40%, a conventional support scheme is adopted; when the decrease in stress concentration factor is 20% to 30% and the reduction in the range of high stress area is 30% to 40%, a reinforced support scheme is adopted; when the decrease in stress concentration factor is less than 20% or the reduction in the range of high stress area is less than 30%, a strengthened support scheme is adopted.

8. The method for controlling mining-induced stress during the final stage of shallow-buried fully mechanized mining as described in claim 1, characterized in that, In step seven, a graded speed reduction method is adopted when resuming advance: after resuming advance, the coal mining machine advances at 50% of the normal speed, and pauses to observe every 2m; when it advances to 5m from the main retreat channel, the speed is reduced to 30% of the normal speed, and pauses every 1m; when it advances to 2m from the main retreat channel, the coal mining machine is stopped, and the breakthrough is completed manually.

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