A determination method for coupled support of ultra-advanced extraction roadway

Abstract

The present invention provides a determination method for the coupled support of the ultra-advanced extraction roadway, belonging to the field of the coupled support of the extraction roadway. Specifically, it includes obtaining the rock mechanics parameters of the roadway to be measured, the active support parameters of the roadway to be measured, the passive support parameters of the roadway to be measured, and the on-site monitoring data; determining the coupling of the main and passive support strengths of the ultra-advanced roadway; determining that the safety factor of the combined main and passive support is within a reasonable range; determining the initial support force of the unit support; determining the coupling of the main and passive support deformations of the ultra-advanced roadway; determining that the unqualified rate of the bolt is within a reasonable range; determining that the subsidence of the live column of the unit support does not exceed the limit; determining that the bulging of the sidewall will not affect the unit support to maintain a stable working posture; observing the coupled support situation, and finally determining whether the coupled support state is reached. The present invention establishes a dual-mode determination mechanism of strength coupling and deformation coupling, and provides a coupling control method, effectively ensuring the strength stability and deformation stability of the extraction roadway.

Description

Technical Field

[0001] The present invention relates to the field of combined support for mining roadways, and specifically includes a method for determining the combined support for the ultra-advanced section of mining roadways. Background Art

[0002] The stability of mining roadways is the key to the safe and efficient mining of coal mines. With the advancement of deep coal resource mining, the surrounding rock stress borne by the roadways increases significantly. Especially in the area ahead of the working face affected by mining, the stress is highly concentrated, far exceeding the original rock stress level. To ensure the stability of mining roadways, reasonable advanced support measures must be adopted to effectively disperse stress, control the deformation of the surrounding rock, and improve the reliability and long-term bearing capacity of the support system.

[0003] The advanced support system for mining roadways is composed of two support forms: active support and passive support. Active support applies pre-tightening force to support materials such as bolts and cables to actively resist the surrounding rock load; passive support is that the unit support passively bears external loads relying on its own strength and stiffness. The combined active and passive support organically combines active support and passive support, comprehensively utilizes the high-strength stability of active support and the flexible adjustment ability of passive support, realizes the coupling of the surrounding rock of the mining roadway, the active support system, and the passive support system in terms of strength and deformation, restricts the harmful deformation damage of the roadway and the support structure, and achieves the goals of integrated support, uniform load, and stable mining roadway.

[0004] Although existing combined support patents and technologies have certain advantages, there are still the following problems: ① Most research focuses on the research and development of new support materials and the innovation of support means, and there is little research on the determination method and implementation process of the coupling state among the active support system mainly composed of bolt support, the passive support system of unit support, and the surrounding rock of the roadway (such as the Chinese patent with the publication number CN216767407U); ② Existing technologies often simply improve the strength stability or deformation stability of the surrounding rock of the roadway through combined support, and do not consider the strength stability and deformation stability as two essential conditions for realizing the overall stability of the roadway and make overall consideration (such as the Chinese patent with the publication number CN114165269B); ③ Some technologies only target specific geological conditions or disaster phenomena (such as the Chinese patent with the publication number CN222457475U), lack universality, and are difficult to be widely applied to other complex mining environments. Based on this, there is an urgent need for a method for determining the combined support for the ultra-advanced section of mining roadways to solve the above problems. Summary of the Invention

[0005] Aiming at the problems existing in the prior art, the present invention provides a method for determining the combined support for the ultra-advanced section of mining roadways.

[0006] The present invention adopts the following technical solution: A method for determining the coupled support of the ultra-advanced mining roadway, comprising the following steps:

[0007] S1: Obtain the rock mechanics parameters of the roadway to be measured, the active support parameters of the roadway to be measured, the passive support parameters of the roadway to be measured, and the on-site monitoring data through on-site investigation.

[0008] S2: Determine the coupling of the active and passive support strengths of the ultra-advanced roadway.

[0009] S2.1: Determine the safety factor of the combined active and passive support within a reasonable range.

[0010] S2.2: Determine the initial support force N of the unit support so that the passive support does not damage the roadway roof.

[0011] After meeting the requirements of S2.1 and S2.2, execute S3.

[0012] S3: Determine the coupling of the active and passive support deformations of the ultra-advanced roadway.

[0013] S3.1: Determine that the unqualified rate of the bolt extension is within a reasonable range.

[0014] S3.2: Determine that the subsidence of the live column of the unit support does not exceed the limit.

[0015] S3.3: Determine that the bulging of the sidewall does not affect the unit support to maintain a stable working posture.

[0016] After meeting the requirements of S3.1, S3.2 and S3.3, execute S4.

[0017] S4: Conduct a comprehensive observation of the coupled support situation and finally determine whether the coupled support state is achieved.

[0018] Preferably, the rock mechanics parameters of the roadway to be measured in S1 include the lithology and structure of the roadway roof, the lithology and structure of the rocks on the roadway sidewall. The active support parameters of the roadway to be measured include the specification parameters of bolts and cables, the layout spacing of bolts and cables, and the layout quantity of bolts and cables. The passive support parameters of the roadway to be measured include the specification parameters of the unit support, the layout spacing of the unit support, and the layout quantity of the unit support. The on-site monitoring data includes the deformation observation data of the roadway surrounding rock, the bolt extension length data, and the subsidence of the live column of the unit support.

[0019] Preferably, the safety factor of the combined active and passive support in S2.1 is:

[0020] .

[0021] In the formula: D is the layout spacing of the unit support, m; B is the roadway section width, m; d gis the bolt row spacing, m; d s is the cable bolt row spacing, m; n g is the number of single-row bolts, pieces; n s is the number of single-row cable bolts, pieces; F g is the designed anchoring force of a single bolt, kN; F s is the designed anchoring force of a single cable bolt, kN; L is the effective anchoring length of the bolt, m; γ is the bulk density of the rock layer where the bolt is anchored, kN / m 3 ; F is the periodic breaking step distance of the main roof, m; K q is the load distribution coefficient, K q = D / 0.5F; m E is the thickness of the main roof, m; m Z is the thickness of the immediate roof, m; γ E is the average bulk density of the rock layer of the main roof, kN / m 3 ; γ Z is the average bulk density of the rock layer of the immediate roof, kN / m 3 ; B0 is the bearing width of the roadway, m; is the roof cutting efficiency, ρ is the density of the rock medium, kg / m 3 ; C P is the wave velocity of the P wave, m / s; ν P ´ is the peak vibration velocity of the particle caused by the P wave; is the number of unit supports within the roof control area of the unit support, is the rated working resistance of the unit support.

[0022] If K l ≥ 2, then the safety factor of the combined active and passive support is within a reasonable range.

[0023] Preferably, S2.2 specifically includes:

[0024] Determine the initial support force N of the unit support so that the bearing pressure on the roadway roof is not greater than the compressive strength σ of the roadway roof rock c .

[0025] The initial support force N of the unit support satisfies:

[0026] .

[0027] In the formula: N is the initial support force of the unit support, kN; A is the effective contact area between the top beam of the unit support and the roof, m 2 ; σ c is the compressive strength of the roadway roof rock, MPa; k is the bearing pressure safety factor.

[0028] Preferably, S3.1 specifically includes: The actual elongation ΔL of the bolt satisfies:

[0029] 。

[0030] If the elongation of the bolt meets the requirements, the bolt elongation is qualified; otherwise, it is unqualified.

[0031] In the formula: σ y is the yield strength of the bolt material, in MPa; E is the elastic modulus of the bolt material, in MPa; L is the effective anchorage length of the bolt, in m.

[0032] The unqualified rate of bolt elongation refers to the proportion of the number of bolts with unqualified bolt elongation in the total number of bolts. If the unqualified rate of bolt elongation is not greater than 3%, the unqualified rate of bolt elongation is within a reasonable range.

[0033] Preferably, S3.2 specifically includes: the ultimate shrinkage amount Δh of the moving support column of the unit support max is:

[0034] Δh max =H max -H min 。

[0035] In the formula: H max is the maximum support height of the unit support, in m; H min is the minimum support height of the unit support, in m.

[0036] The actual shrinkage amount Δh of the moving support column of the unit support is equal to the shrinkage amount of the roadway roof and floor.

[0037] Determine that the actual shrinkage amount Δh of the moving support column of the unit support does not exceed the ultimate shrinkage amount Δh of the moving support column max That is:

[0038] Δh max＞ μΔh.

[0039] In the formula, μ is the safety factor of the shrinkage amount.

[0040] Preferably, S3.3 specifically includes: calculating the overturning moment of the unit support. The overturning moment of the unit support is:

[0041] 。

[0042] M G is the self-weight overturning moment, which is the overturning moment generated by the self-weight G of the unit support on the overturning edge; M N1 is the self-supporting overturning moment of the unit support roof, which is the moment generated by the vertical force on the top beam of the unit support on the overturning edge; M N2 is the self-supporting overturning moment of the unit support base, which is the moment generated by the vertical force on the base of the unit support on the overturning edge; M f1 is the friction overturning moment of the unit support roof, which is the moment generated by the friction force on the upper surface of the top beam of the unit support on the overturning edge; M f2$M_{f}$ is the frictional overturning moment of the unit support base, which is the moment generated by the frictional force on the lower surface of the unit support base acting on the overturning edge; $M_{e}$ is the external force overturning moment, which is the moment generated by the horizontal force exerted by the bulge of the roadway rib on the unit support acting on the overturning edge.

[0043] It is determined that the overturning moment $M$ of the unit support is greater than 200.

[0044] Preferably, S4 specifically includes:

[0045] S4.1: Observe the deformation of the surrounding rock of the roadway: One week after the implementation of the above S1 to S3 coupling steps, observe the deformation of the surrounding rock of the roadway for a continuous week. Check the surrounding rock at the top, rib, and bottom of the roadway ahead of the advanced section every day, and pay attention to whether there are obvious roof collapses, rib bulges, and floor bulges; at the location where the deformation of the surrounding rock of the roadway is the largest, if the cumulative deformation within a week does not exceed 15 mm, it is considered that the deformation of the surrounding rock of the roadway is normal.

[0046] S4.2: Observe the working condition of the active support system: One week after the implementation of the above S1 to S3 coupling steps, observe the deformation of the active support for a continuous week. Check the state of the active support system every day, and pay attention to the elongation and anchoring condition of the bolts; within a cumulative week, if the number of bolts whose elongation length exceeds their maximum elongation length does not exceed two, and there is no bolt fracture, it is considered that the active support system is working normally.

[0047] S4.3: Observe the working condition of the passive support system: One week after the implementation of the above S1 to S3 coupling steps, observe the deformation of the passive support for a continuous week. Check the state of the active support system every day, and pay attention to the height change and stability of the unit support; within a cumulative week, if there is no support being crushed and no instability and overturning of the unit support occur, it is considered that the passive support system is working normally.

[0048] The beneficial effects of the present invention are:

[0049] 1. The present invention provides a determination method for the coupled support of the advanced section of the mining roadway, which for the first time makes an overall coupled determination of the active support, passive support, and the surrounding rock of the roadway, and establishes a dual-mode determination mechanism of strength coupling and deformation coupling. This mechanism breaks through the limitation of the traditional determination method that only focuses on strength or deformation, and ensures the scientificity and comprehensiveness of the determination result.

[0050] 2. The determination method for the coupled support of the advanced section of the mining roadway provided by the present invention comprehensively uses the methods of coefficient correction and theoretical calculation based on on-site data, and provides a scientific and quantitative basis for coupled determination. Compared with the traditional empirical discrimination method, this method has stronger operability and effectively improves the safety of the overall support of the mining roadway in the mine.

[0051] 3. The determination method of coupled support for the ultra - front extraction roadway provided by the present invention monitors the deformation of the bolt group in the roadway to be measured in real time through an optical fiber synchronous deformation composite bolt. This device has the ability of full - distribution and high - precision monitoring, can adapt to the complex underground environment and achieve long - term stable monitoring. An evaluation method for the overturning instability of the unit support is also established to quantitatively evaluate the instability state of the unit support, which can guide the scientific use of the unit support.

[0052] 4. The coupled determination method provided by the present invention realizes the closed - loop management of "evaluation - early warning - regulation". By accurately evaluating the coupled state of the support system, differential regulation strategies are accurately formulated. Further, this process method can be combined with an automated monitoring system and a remote control platform to form a more intelligent support management system, further improving the mine safety, production efficiency and economic benefits. BRIEF DESCRIPTION OF THE DRAWINGS

[0053] Figure 1 It is a schematic diagram of the data acquisition system for the actual extension length of the bolt.

[0054] Figure 2 It is a schematic diagram of the optical fiber synchronous deformation composite bolt.

[0055] Figure 3 It is a flow chart of the present invention.

[0056] Figure 4 It is a schematic diagram of the overturning instability of the unit support.

[0057] Figure 5 It is a force analysis diagram when the unit support undergoes overturning instability.

[0058] Figure 6 It is a force analysis diagram of the unit support in Embodiment 1. DETAILED DESCRIPTION OF THE INVENTION

[0059] The following further describes the specific embodiments of the present invention in conjunction with the drawings and specific embodiments: Figures 1 to 6 , a determination method of coupled support for the ultra - front extraction roadway, includes the following steps:

[0060] S1: Obtain the rock mechanics parameters of the roadway to be measured, the active support parameters of the roadway to be measured, the passive support parameters of the roadway to be measured and the on - site monitoring data through on - site investigation.

[0061] The rock mechanics parameters of the roadway to be measured include the lithology and structure of the roadway roof, and the lithology and structure of the rock on the roadway sides. The active support parameters of the roadway to be measured include the specification parameters of bolts and cables, the layout spacing of bolts and cables, and the layout quantity of bolts and cables. The passive support parameters of the roadway to be measured include the specification parameters of unit supports, the layout spacing of unit supports, and the layout quantity of unit supports. The on-site monitoring data includes the deformation observation data of the roadway surrounding rock, the bolt extension length data, and the subsidence amount of the unit support's movable column.

[0062] S2: Coupling determination of the active and passive support strengths of the ultra-advanced roadway.

[0063] S2.1: Determine the safety factor of the combined active and passive support Within a reasonable range.

[0064] Calculate the maximum support strength that the active bolt support can provide, the maximum passive support strength that the unit support can provide, and the roadway roof load. Determine whether the maximum support strength that the combined active and passive support can provide can resist the maximum roof load, and then determine the combined active and passive safety factor K l Within a reasonable range.

[0065] Combined active and passive safety factor K l The calculation principle is: .

[0066] Furthermore, we get: .

[0067] In the formula: p l is the maximum support strength that the combined active and passive support system can provide, MPa; S k is the roof control area of the unit support, m 2 ; P Z is the active support strength, MPa; P B is the passive support strength, MPa; Q c is the maximum surrounding rock load, kN; Q1 is the loosening load of the roof surrounding rock, kN; Q2 is the mining-induced load, kN; Q3 is the impact load, kN.

[0068] The final safety factor of the combined active and passive support is:

[0069] ;

[0070] In the formula: D is the layout spacing of the unit support, m; B is the roadway section width, m; d g is the bolt row spacing, m; d s is the cable row spacing, m; n g is the number of bolts in a single row, roots; n s is the number of cables in a single row, roots; F gAnchoring force designed for a single bolt, kN; F s Anchoring force designed for a single cable bolt, kN; L is the effective anchoring length of the bolt, m; γ is the unit weight of the rock layer where the bolt is anchored, kN / m 3 .

[0071] F is the periodic fracture step distance of the main roof, m; K q is the load distribution coefficient, K q = D / 0.5F; m E is the thickness of the main roof, m; m Z is the thickness of the immediate roof, m; γ E is the average unit weight of the rock layer of the main roof, kN / m 3 ; γ Z is the average unit weight of the rock layer of the immediate roof, kN / m 3 ; B0 is the load-bearing width of the roadway, m; is the roof cutting efficiency, ρ is the density of the rock medium, kg / m 3 ; C P is the wave velocity of the P-wave, m / s; ν P ´ is the peak particle vibration velocity caused by the P-wave; is the number of unit supports within the roof control area of the unit support, is the rated working resistance of the unit support.

[0072] Judgment evaluation and treatment measures for the safety factor K of the combined active and passive support are shown in Table 1. l are as shown in Table 1.

[0073] Table 1

[0074]

[0075] Reinforcement support: ① Increase the bolt support density, that is, add bolts; ② Increase the strength of the passive support, that is, reduce the layout spacing of the unit supports.

[0076] If K l ≥ 2, then the safety factor of the combined active and passive support is reasonable. If the safety factor of the combined active and passive support is too small, then adjust the safety factor of the combined active and passive support within a reasonable range through reinforcement support.

[0077] S2.2: Determine the initial support force N of the unit support so that the passive support does not damage the roadway roof.

[0078] Setting too large an initial support force N during the initial support stage of the unit support will cause the top beam of the unit support to damage the rock layer of the roadway roof, resulting in the phenomenon of the unit support "drilling into the roof". To prevent the occurrence of the drilling-into-the-roof phenomenon, a reasonable initial support force N of the unit support should be determined so that the specific pressure on the roadway roof is not greater than the compressive strength σ of the roadway roof rockc 。

[0079] The specific pressure on the roadway roof is the ratio of the initial support force of the unit support to the area of the unit support roof beam. It is necessary to ensure that the specific pressure on the roadway roof is not greater than the compressive strength σc of the roof rock, that is, the initial support force N of the unit support should satisfy:

[0080] 。

[0081] In the formula: N is the initial support force of the unit support, kN; A is the area of the unit support roof beam, m 2 ; σ c is the compressive strength of the roadway roof rock, MPa; k is the safety factor of specific pressure, usually taken as 2 - 3.

[0082] If N ≥ (Aσ c ) / k, then it is necessary to reduce the initial support force of the unit support until N < (Aσ c ) / k. The prerequisite for reducing the initial support force of the unit support is to strictly comply with the coal mine management regulations and meet the initial support force requirements.

[0083] S2.3: After meeting the requirements of S2.1 and S2.2, execute S3.

[0084] That is, the coupling determination of the active and passive support strengths of the pre - advanced section of the roadway needs to satisfy K l ≥ 2, 。

[0085] S3: Coupling determination of the active and passive support deformations of the pre - advanced section of the roadway.

[0086] S3.1: Determine whether the elongation rate of the bolt is within a reasonable range.

[0087] The maximum elongation length of the bolt refers to the maximum axial deformation that the bolt can produce before it enters the plastic stage. When the bolt deformation enters the plastic stage, it means that the bolt deformation has exceeded the elastic limit, and the bolt no longer fully returns to its original shape. Even after the load is removed, it will leave some permanent deformation. This indicates that the bearing capacity of the bolt has decreased, or the bolt is approaching its failure limit. For the support system, the bolt entering the plastic stage may affect the stability of the support. Especially when the bolt continues to bear excessive loads, it may lead to support failure or further deformation of the surrounding rock mass.

[0088] Therefore, when designing the support system, it is necessary to ensure that the bolt deformation does not enter the plastic stage, or take appropriate measures when necessary to ensure that the support system can still effectively provide support and prevent rock mass collapse or support failure.

[0089] Determine the reasonable range of the actual elongation length ΔL of the bolt, that is, the actual elongation length ΔL of the bolt satisfies:

[0090] .

[0091] In the formula: σ y is the yield strength of the bolt material, in MPa; E is the elastic modulus of the bolt material, in MPa; L is the effective anchorage length of the bolt, in m.

[0092] If the above conditions are met, it is determined that the bolt extension is qualified; otherwise, it is determined that the bolt extension is unqualified.

[0093] It should be noted that: in the mining roadway, bolts do not exist individually. The bolt group plays a supporting role together. The failure or invalidation of a single bolt usually does not cause the overall support failure of the roadway or the loss of control of the surrounding rock. Therefore, the object of monitoring in S3.1 is the bolt group in the mining roadway. The unqualified rate of bolt extension refers to the proportion of the number of bolts with unqualified bolt extension to the total number of bolts. If the unqualified rate of bolt extension is not greater than 3%, the unqualified rate of bolt extension is within a reasonable range.

[0094] If the unqualified rate of bolt extension exceeds 3%, a bolt shall be additionally drilled beside the bolt with unqualified extension length.

[0095] In the mining roadway, the number of bolts is huge. If the actual extension length ΔL of each bolt is monitored one by one, the workload is extremely large and the monitoring cost is extremely high. To solve the above problems, the present invention designs a fiber optic synchronous deformation composite bolt, as Figure 1 and 2 , and the fiber optic synchronous deformation composite bolt is used to monitor the actual extension length ΔL of the bolt group in real time.

[0096] Production and assembly steps of the fiber optic synchronous deformation composite bolt:

[0097] Step 1: Bolt grooving: Along the axial direction of bolt 1, parallel to the bolt central axis, a micro-groove 2 is opened on the bolt surface. The depth of micro-groove 2 ≤ 3% of the bolt diameter, and the width of micro-groove 2 is 1 - 1.5 mm.

[0098] Step 2: Fiber optic embedding: Select a Brillouin distributed fiber optic sensor 3 with a fiber diameter of 100 - 250 μm; lay the fiber optic in the micro-groove 2 to ensure that the fiber optic is taut and has no bending; reserve a 5 cm fiber optic buffer zone at the anchorage end of the bolt, and the fiber optic end is enclosed and protected with a protective cover 4; reserve the installation position of the fiber optic connector 8 at the exposed end of the bolt and connect it to the data acquisition system.

[0099] Step 3: Backfill the micro-groove with a high-strength filling material: Use a carbon fiber reinforced epoxy resin material 5 to fill the micro-groove to ensure that the fiber optic is tightly attached to the bolt; after the filling material is cured, polish the bolt surface to ensure that the bolt surface is flat and avoid the protrusion of the filling area affecting the anchorage of the bolt to the surrounding rock.

[0100] Step 4: Bolt installation: Use a bolt drilling rig to drill holes at the calibrated hole positions on the roadway roof 12; after drilling, insert 1 quick-setting anchoring agent and 1 medium-setting anchoring agent in sequence; connect the bolt to the bolt agitator, insert the bolt into the hole and slowly agitate the anchoring agent, with the agitation time being 25 - 35 seconds and the gel time being 90 seconds; after waiting for 8 minutes, use the bolt drilling rig to tighten the nut 7 to press the bolt tray 6 against the rock face; use a pneumatic hammer or torque wrench to tighten the bolt a second time to reach the designed pre-tightening torque.

[0101] The data acquisition system for the actual extension length of the bolt is as Figure 1 , and its assembly sequence:

[0102] Optical fiber synchronous deformation composite bolt → Reserved optical fiber end at the exposed end of the bolt → Optical fiber connector 8 → Brillouin optical time domain reflectometer device 9 → Data acquisition card 10 → Computer 11.

[0103] Usage method of the data acquisition system for the actual extension length of the bolt:

[0104] ① After all devices are powered on, the Brillouin optical time domain reflectometer 9 scatters light sources into the Brillouin distributed optical fiber sensor inside the bolt.

[0105] ② The Brillouin distributed optical fiber sensor 3 measures the change in Brillouin frequency shift and calculates the strain value of the optical fiber.

[0106] ③ The Brillouin optical time domain reflectometer 9 receives and processes the signals returned by the optical fiber sensor.

[0107] ④ The data acquisition card 10 connects the computer and the Brillouin optical time domain reflectometer 9 to convert the optical fiber signal into digital information.

[0108] ⑤ Through the computer 11 software, analyze the actual deformation amount of the bolt, and output and display the result of the actual deformation amount of the bolt.

[0109] S3.2: Ensure that the telescopic amount of the telescopic column of the unit support does not exceed the limit.

[0110] With the manifestation of mine pressure, the roadway section converges and the roof-to-floor convergence amount increases. If the roadway convergence amount exceeds the limit of the telescopic amount of the telescopic column of the unit support, the "telescopic column of the unit support will be crushed", resulting in damage to the support equipment and seriously endangering the safety of life and property underground. The necessary condition to avoid the telescopic column of the unit support from being crushed is that the actual telescopic amount Δh of the telescopic column of the unit support cannot exceed the limit telescopic amount Δh of the telescopic column max , and the calculation process is as follows:

[0111] The limit telescopic amount Δh of the telescopic column of the unit support max is:

[0112] Δh max =H max -H min .

[0113] Where: H max is the maximum support height of the unit support, in m; H min is the minimum support height of the unit support, in m.

[0114] The actual retraction amount Δh of the movable column of the unit support is determined by on-site monitoring. The actual retraction amount Δh of the movable column of the unit support is equal to the shrinkage amount of the roadway roof and floor.

[0115] Determine that the actual retraction amount Δh of the movable column of the unit support does not exceed the limit retraction amount Δh of the movable column max That is:

[0116] Δh max＞ ≤ μΔh.

[0117] Where μ is the safety factor of the retraction amount, taking 1.2 - 1.5.

[0118] If Δh max≤ > μΔh, then strengthen the support density of the unit support so that Δh max＞ ≤ μΔh. Strengthening the support density of the unit support means increasing the number of unit supports and reducing the layout spacing of unit supports.

[0119] S3.3: Determine that the bulging of the sidewall will not affect the unit support to maintain a stable working posture.

[0120] As the coal mining face advances, the stress of the roadway surrounding rock redistributes, and the stress borne by the sidewall increases. Also, due to reasons such as insufficient support strength, the surrounding rock of the sidewall gradually bulges and squeezes the unit support, and it is easy to cause the unit support to gradually lose stability, and finally lead to the unit support overturning and losing stability into the roadway. The fundamental reason for the unit support to overturn and lose stability is that the surrounding rock of the sidewall gradually bulges and squeezes the unit support, and the horizontal force exerted by the sidewall on the unit support destroys the balanced working state of the unit support.

[0121] The risk of the unit support tipping and losing stability can be judged by the overturning moment M of the unit support. The overturning moment of the unit support refers to the product of the overturning load acting on the unit support and the overturning force arm (the distance from the overturning load to the overturning edge). According to different overturning loads, the overturning moment of the unit support is divided into the self-weight overturning moment of the unit support, the support overturning moment of the unit support, and the external force overturning moment of the unit support. The self-weight overturning moment of the unit support refers to the overturning moment M G generated by the self-weight G of the unit support on the overturning edge. The support overturning moment of the unit support refers to the moment M N1 generated by the vertical force N1 on the top beam of the unit support, the vertical force N2 on the base of the unit support, the friction force f1 on the upper surface of the top beam of the unit support, and the friction force f2 on the lower surface of the base of the unit support on the overturning edge after the unit support exerts a support resistance on the roadway roof and floor. N2 、M f1 、Mf2 The external overturning moment refers to the horizontal force F exerted by the bulging of the roadway sidewall on the unit support X that generates an overturning moment M on the overturning edge Fx The combined moment refers to the vector sum of the above moments. The unit support maintains stability through the action of the combined moment. To prevent the unit support from tipping over and ensure stability, the key is to ensure that the combined overturning moment is greater than 0 (clockwise moment is specified as positive).

[0122] Calculate the combined overturning moment of the unit support. The combined overturning moment of the unit support is as follows:

[0123] 。

[0124] The overturning force arm refers to the distance from the overturning load to the overturning edge. As Figure 5 shown. H G is half of the width of the bottom plate of the unit support, m; H N1 is half of the width of the top beam of the unit support, m; N2 is half of the width of the base of the unit support, m; H f1 is the support height of the unit support (roadway height), m; H f2 is 0 m; H Fx is half of the support height of the unit support (roadway height), m.

[0125] M G is the overturning moment generated by the self-weight G of the unit support on the overturning edge, M N1 is the moment generated by the vertical force on the top beam of the unit support on the overturning edge, M N2 is the moment generated by the vertical force on the base of the unit support on the overturning edge, M f1 is the moment generated by the friction force on the upper surface of the top beam of the unit support on the overturning edge, M f2 is the moment generated by the friction force on the lower surface of the base of the unit support on the overturning edge; is the moment generated by the horizontal force exerted by the bulging of the roadway sidewall on the unit support on the overturning edge.

[0126] Determine whether the overturning moment M of the unit support is greater than 200

[0127] If M is less than or equal to 200, the unit support lowers the support to adjust its posture, and repairs the surrounding rock of the sidewall so that it no longer protrudes and squeezes the unit support.

[0128] S3.4: After meeting the requirements of S3.1, S3.2, and S3.3, execute S4.

[0129] S4: Conduct a comprehensive observation on the coupled support situation to finally determine whether the coupled support state is reached.

[0130] S4.1: Observe the deformation of the surrounding rock of the roadway: One week after the implementation of the above S1 to S3 coupling steps, observe the deformation of the surrounding rock of the roadway for one continuous week. Check the surrounding rock at the top, sides, and bottom of the advanced section of the roadway every day, and pay attention to whether there are obvious roof collapses, side bulges, and floor bulges. If the cumulative deformation within a week at the location with the largest deformation of the roadway surrounding rock does not exceed 15 mm, it is considered that the deformation of the roadway surrounding rock is normal.

[0131] S4.2: Observe the working condition of the active support system: One week after the implementation of the above S1 to S3 coupling steps, observe the deformation of the active support for one continuous week. Check the status of the active support system every day, and pay attention to the elongation and anchoring conditions of the bolts. If the number of bolts with an elongation length exceeding their maximum elongation length within one week does not exceed two, and there is no bolt fracture, it is considered that the active support system is working normally.

[0132] S4.3: Observe the working condition of the passive support system: One week after the implementation of the above S1 to S3 coupling steps, observe the deformation of the passive support for one continuous week. Check the status of the active support system every day, and pay attention to the height change and stability of the unit support. If there is no support crushing or unit support instability and overturning within one week, it is considered that the passive support system is working normally.

[0133] If all the above observations are normal, it is determined that the coupled support state is reached.

[0134] Example 1. The following is a verification and explanation of the above coupling determination method for the support of the advanced section of the mining roadway through a specific construction site case of the track gateway of the 2309 working face in a certain coal mine:

[0135] S1: Obtain the rock mechanics parameters of the roadway to be measured, the active support parameters of the roadway to be measured, the passive support parameters of the roadway to be measured, and the on-site monitoring data through on-site investigation.

[0136] Basic parameters and abutment pressure conditions of the roadway strata:

[0137] The track roadway is driven along the goaf, and the width B of the coal pillar m = 4 m, arranged along the coal seam floor with roof coal supported. The cross-section is rectangular, with a net width B of 5.6 m and a net height of 3.8 m. The elastic-plastic demarcation length B of the solid coal side s = 3.15 m; the bulk density γ of the solid coal: 14 kN / m 3 ; the thickness m of the immediate roof Z = 4.48 m, and the average bulk density γ of the immediate roof Z = 23 kN / m 3 ; the thickness m of the main roof E = 11.51 m; the average bulk density γ of the main roof E = 23 kN / m 3 .

[0138] The periodic fracture step distance of the main roof F = 13 m; the medium density ρ = 23 kg / m 3 , the P-wave velocity C P = 5.8×10 3 m / s, and the vibration velocity v´ P = 1.37 m / s.

[0139] Active support conditions:

[0140] The form and specifications of the bolt: Φ22×2800 mm left-handed ribless threaded steel bolt, the designed anchoring force of a single bolt F g = 190 kN, and the yield strength σ of the bolt material y = 500 MPa; the moment of inertia I of the bolt cross-section = 11493.185 mm 4 .

[0141] Bolt layout: The bolt spacing is 0.9 m, the row spacing d g = 0.9 m, the number of single-row bolts n g = 7, and L is the effective anchoring length of the bolt, 2.65 m; the unit weight γ of the rock layer where the bolt is anchored = 13 kN / m 3 .

[0142] The form and specifications of the cable bolt: 1×19S structural steel strand cable bolt with Φ21.8 mm×8200 mm. The number of single-row cable bolts n s = 3, and the designed anchoring force of a single cable bolt F s = 553 kN.

[0143] Cable bolt layout: The spacing is 1.8 m, and the row spacing d s = 1.8 m.

[0144] Passive support conditions: Use the ZQ4000 / 20.6 / 45 type unit support to support the roof, with a rated working resistance N d = 4000 kN; the minimum support height H min = 2.06 m; the maximum support height H max = 4.5 m; the top beam size: 1600 mm×780 mm, the base size: 1812 mm×940 mm. The effective contact area A between the top beam of the unit support and the roof = 1.248 m 2 ; the unit support is arranged in double rows n d = 2 frames, the layout spacing D of the unit support = 6 m, and the initial support force is set to 600 kN during installation.

[0145] S2: Coupling judgment of the main and passive support strengths in the ultra-advanced roadway section.

[0146] S2.1: Calculation of the safety factor K of the main and passive combined support l :

[0147] =2.4.

[0148] In the formula: In the formula, ξ is the top breaking efficiency, which is taken as 1.0; B0 is the bearing width of the roadway, B0=B+B m +B S =12.75m; K q is the load distribution factor, K q =D / 0.5F=0.92.

[0149] K l =2.4>2,K l Within a reasonable range.

[0150] S2.2: Determine the initial support force N of the unit support so that the passive support will not damage the tunnel roof.

[0151] =6364.8kN.

[0152] Where: N is the initial support force of the unit support, which is 600kN on site; A is the area of the top beam of the unit support, 1.248m 2 ; σ c is the compressive strength of the tunnel roof rock, 15.3MPa; k is the specific pressure safety factor, which is 3.

[0153] The initial support force set on site is 600N, which is much smaller than 6364.8kN, so the initial support force will not destroy the integrity of the tunnel roof.

[0154] In summary, the track chute of the 2309 working face meets the strength coupling standard and can proceed to the next step of deformation coupling determination.

[0155] S3: Determination of coupling deformation of active and passive support in the leading tunnel.

[0156] S3.1: Determine that the anchor extension failure rate is within a reasonable range.

[0157] Determine the reasonable range of the actual extension length ΔL of the anchor rod, that is:

[0158] =1.325cm.

[0159] Where: yield strength of anchor material σ y =500MPa; effective anchoring length of anchor rod L = 2.65m; elastic modulus of anchor rod material E = 1×10 5 MPa.

[0160] Feedback from on-site monitoring showed that the actual extension length of all anchor rods did not exceed 1.325cm, that is, the anchor rod extension failure rate was 0%, and subsequent monitoring was normal.

[0161] S3.2: Determine that the retraction amount of the moving support column of the unit support does not exceed the limit.

[0162] The limit retraction amount Δh of the moving support column of the unit support max =H max -H min = 2.44 m.

[0163] On-site monitoring data shows that the maximum shrinkage amount of the roof and floor of the track gateway in the 2309 working face is 0.2 m, that is, the actual retraction amount Δh of the moving support column of the unit support is 0.2 m. Obviously:

[0164] Δh max＞ μΔh = 0.3 m.

[0165] Note: In the formula, μ is the safety factor of the retraction amount, taking 1.5.

[0166] The actual height of the roadway section needs to be greater than the minimum height of the unit support, and the unit support in the 2309 track gateway will not be crushed. Conduct subsequent normal monitoring.

[0167] S3.3: Determine that the bulging of the rib will not affect the unit support to maintain a stable working posture.

[0168] The steady-state analysis method of the unit support is as follows:

[0169] To calculate the overturning moment of the unit support, take the unit support as the research object and conduct a force analysis as Figure 6 :

[0170] = 4003.38 > 200.

[0171] Among them: The self-weight overturning moment M G : It refers to the overturning moment generated by the self-weight G of the unit support on the overturning edge. Among them, the gravity G = 54 kN, and the overturning force arm H G = 0.47 m.

[0172] The self-supporting overturning moment M of the roof of the unit support N1 : It refers to the moment generated by the vertical force on the top beam of the unit support on the overturning edge after the unit support exerts a support resistance on the roof and floor of the roadway. Among them, N1 = 4000 kN, and the overturning force arm H N1 = 0.47 m.

[0173] The self-supporting overturning moment M of the base of the unit support N2 : It refers to the moment generated by the vertical force on the base of the unit support on the overturning edge after the unit support exerts a support resistance on the roof and floor of the roadway. Among them, N2 = 4000 kN, and the overturning force arm H N2 = 0.47 m.

[0174] The frictional overturning moment M of the roof of the unit support f1: It refers to the moment generated by the frictional force on the upper surface of the unit support beam on the tipping edge after the unit support exerts a support resistance on the roof and floor of the roadway, where f1 = μN1 = 1200 kN and the tipping force arm H f1 = 3.6 m.

[0175] Tipping moment M of the unit support base friction f2 : It refers to the moment generated by the frictional force on the lower surface of the unit support base on the tipping edge after the unit support exerts a support resistance on the roof and floor of the roadway, where f2 = μN2 = 1200 kN and the tipping force arm H f1 = 0 m.

[0176] External force tipping moment M Fx : It refers to the moment generated by the horizontal force exerted by the bulging of the roadway sidewall on the unit support on the tipping edge, where F X = 190 kN and the tipping force arm H Fx = 1.8 m.

[0177] ① The bending moment calculation is positive in the clockwise direction; when the unit support is unstable and tips inward into the roadway, the tipping edge is the inner side edge of the unit support base in the roadway, and the tipping point is the lowest point of the inner side edge of the unit support base in the roadway.

[0178] ② The weight of the unit support is 5400 kg, the gravity G = 54 kN, and the gravitational acceleration g is taken as 10 m / s 2 ; The unit support beam and base receive reaction forces from the surrounding rock. Considering the most dangerous situation, the output working resistance of the unit support is 4000 kN, that is, N1 = N2 = 4000 kN; there is a frictional force f1 between the unit support beam and the roof rock, f1 = μN1 = 1200 kN; there is a frictional force f2 between the unit support base and the floor rock, f2 = μN2 = 2000 kN, and μ is the friction coefficient taken as 0.5; the horizontal force F exerted by the bulging of the roadway sidewall on the unit support x = 190 kN.

[0179] The tipping moment M of the unit support should be greater than 200. It is judged that the unit support in the track roadway of the 2309 working face can maintain a steady state, is in a safe state, and subsequent normal monitoring is carried out.

[0180] In summary, the track roadway of the 2309 working face meets the deformation coupling standard.

[0181] S4: Conduct comprehensive observations on the coupling support situation and finally determine whether the coupling support state is achieved.

[0182] After one week of implementation of the above S1-S3 coupling steps, the roadway surrounding rock deformation, the working conditions of the active support system, and the working conditions of the passive support system of the track roadway in the 2309 working face were observed for one week. The observation results showed that: ① At the location with the largest deformation of the roadway surrounding rock, the cumulative weekly deformation was 8 mm; ② There was no bolt with an extended length exceeding its maximum extended length, and no bolt fracture occurred; ③ There was no situation where the support was crushed or the unit support was unstable and overturned. Finally, it was determined that the roadway reached the coupled support state.

[0183] Of course, the above description is not a limitation of the present invention, and the present invention is not limited to the above examples. Changes, modifications, additions, or substitutions made by those skilled in the art within the scope of the essence of the present invention should also fall within the protection scope of the present invention.

Claims

1. A determination method for coupling support of ultra - front extraction roadway, characterized in that, It includes the following steps: S1: Conduct on-site investigation to obtain the rock mechanics parameters of the roadway to be measured, the active support parameters of the roadway to be measured, the passive support parameters of the roadway to be measured, and the on-site monitoring data; S2: Determine the coupling of the active and passive support strengths in the ultra-advanced roadway section; S2.1: Determine the safety factor K of the combined active and passive support l Within a reasonable range; The safety factor K of the combined active and passive support l is as follows: Where: D is the layout spacing of unit supports, m; B is the roadway cross-section width, m; d g is the bolt row spacing, m; d s is the cable row spacing, m; n g is the number of bolts in a single row, pieces; n s is the number of cables in a single row, pieces; F g is the designed anchoring force of a single bolt, kN; F s is the designed anchoring force of a single cable, kN; L is the effective anchoring length of the bolt, m; γ is the unit weight of the bolt-anchored rock stratum, kN / m 3 ; F is the periodic breaking step distance of the main roof, m; K q is the load distribution coefficient, K q = D / 0.5F; m E is the thickness of the main roof, m; m Z is the thickness of the immediate roof, m; γ E is the average unit weight of the main roof rock stratum, kN / m 3 ; γ Z is the average unit weight of the immediate roof rock stratum, kN / m 3 ; B0 is the roadway bearing width, m; ξ is the roof-breaking efficiency, ρ is the density of the rock medium, kg / m 3 ; C P is the P-wave velocity, m / s; ν P ′ is the peak vibration velocity of the particle caused by the P-wave; n d is the number of unit supports within the roof control area of the unit support, N d is the rated working resistance of the unit support; If K l ≥ 2, then the safety factor K of the combined active and passive support l is within a reasonable range; S2.2: Determine the initial support force N of the unit support so that the passive support will not damage the roadway roof; S2.3: After meeting the requirements of S2.1 and S2.2, execute S3; S3: Determine the coupling of the active and passive support deformations in the ultra-advanced roadway section; S3.1: Ensure that the unqualified rate of bolt extension is within a reasonable range; S3.2: Ensure that the subsidence of the telescopic column of the unit support does not exceed the limit; S3.3: Ensure that the bulging of the side does not affect the unit support's ability to maintain a stable working posture; S3.4: After meeting the requirements of S3.1, S3.2, and S3.3, execute S4; S4: Conduct comprehensive observation of the coupled support situation and finally determine whether the coupled support state is achieved.

2. The determination method of coupled support for the ultra-advanced mining roadway according to claim 1, characterized in that, The rock mechanics parameters of the roadway to be measured in S1 include the lithology and structure of the roadway roof, and the lithology and structure of the rocks on the sides of the roadway. The active support parameters of the roadway to be measured include the specification parameters of bolts and cables, the layout spacing of bolts and cables, and the number of bolts and cables arranged. The passive support parameters of the roadway to be measured include the specification parameters of the unit support, the layout spacing of the unit support, and the number of unit supports arranged. The on-site monitoring data includes the roadway surrounding rock deformation observation data, the bolt extension length data, and the subsidence of the telescopic column of the unit support.

3. The determination method of coupled support for the ultra-advanced extraction roadway according to claim 1, characterized in that S2.2 specifically includes: Determine the initial support force N of the unit support so that the bearing pressure on the roadway roof is not greater than the compressive strength σ of the roadway roof rock c ; The initial support force N of the unit support satisfies: Where: N is the initial support force of the unit support, kN; A is the effective contact area between the top beam of the unit support and the roof, m 2 ; σ c is the compressive strength of the roadway roof rock, MPa; k is the specific pressure safety factor.

4. The determination method of coupled support for the ultra-advanced extraction roadway according to claim 1, characterized in that S3.1 specifically includes: The actual bolt extension length ΔL satisfies: If so, it is determined that the bolt extension is qualified; otherwise, it is determined that the bolt extension is unqualified; Where: σ y is the yield strength of the bolt material, in MPa; E is the elastic modulus of the bolt material, in MPa; L is the effective anchorage length of the bolt, in m; The unqualified rate of bolt extension refers to the proportion of the number of bolts with unqualified bolt extensions to the total number of bolts. If the unqualified rate of bolt extension is not greater than 3%, the unqualified rate of bolt extension is within a reasonable range.

5. The determination method of coupled support for the ultra-advanced extraction roadway according to claim 1, characterized in that S3.2 specifically includes: The ultimate downward shrinkage Δh of the movable pillar of the unit support max is as follows: Δh max = H max - H min ; Where: H max is the maximum support height of the unit support, m; H min is the minimum support height of the unit support, m; The actual subsidence Δh of the telescopic column of the unit support is equal to the subsidence of the roadway roof and floor; Determine that the actual retraction amount Δh of the unit support's movable column does not exceed the limit retraction amount Δh of the movable column max That is: Δh max＞ μΔh; In the formula, μ is the safety factor for subsidence.

6. The determination method of coupled support for ultra-advanced extraction roadway according to claim 1, characterized in that S3.3 specifically includes: Calculate the overturning moment of the unit support. The overturning moment of the unit support is: M = M G + M N1 + M N2 + M f1 + M f2 + M Fx ; M G is the self-weight overturning moment, which is the overturning moment generated by the self-weight G of the unit support on the overturning edge; M N1 is the self-supporting overturning moment of the unit support roof, which is the moment generated by the vertical force on the top beam of the unit support on the overturning edge; M N2 is the self-supporting overturning moment of the unit support base, which is the moment generated by the vertical force on the unit support base on the overturning edge; M f1 is the frictional force overturning moment of the unit support roof, which is the moment generated by the frictional force on the upper surface of the top beam of the unit support on the overturning edge; M f2 is the frictional force overturning moment of the unit support base, which is the moment generated by the frictional force on the lower surface of the unit support base on the overturning edge; M Fx is the external force overturning moment, which is the moment generated by the horizontal force exerted by the bulging of the roadway side on the unit support on the overturning edge; Ensure that the overturning moment M of the unit support is greater than 200.

7. The determination method of coupled support for the ultra-advanced extraction roadway according to claim 1, characterized in that, S4 specifically includes: S4.1: Observe the deformation of the roadway surrounding rock: One week after the implementation of the coupling steps from S1 to S3 above, conduct a one-week observation of the deformation of the roadway surrounding rock. Check the surrounding rock at the top, sides, and bottom of the ultra-advanced roadway section every day, and pay attention to whether there are obvious roof collapses, side bulges, and floor bulges; at the location with the largest deformation of the roadway surrounding rock, if the cumulative deformation within a week does not exceed 15 mm, it is considered that the deformation of the roadway surrounding rock is normal; S4.2: Observe the working condition of the active support system: One week after the implementation of the coupling steps from S1 to S3 above, conduct a one-week observation of the deformation of the active support. Check the status of the active support system every day and pay attention to the elongation and anchoring conditions of the bolts; within a week, if the number of bolts with an extension length exceeding their maximum extension length does not exceed two and there is no bolt fracture, it is considered that the active support system is working normally; S4.3: Observe the working condition of the passive support system: One week after the implementation of the above S1 to S3 coupling steps, observe the deformation of the passive support for one week continuously, check the status of the active support system every day, and pay attention to the height change and stability of the unit support; If there is no situation where the support is crushed or the unit support is unstable and overturned within one week, it is regarded that the passive support system is working normally.

Images