Determination method for coupling support of over-forepart mining roadway

Abstract

The invention provides a method for determining a coupling support of an over-forepart mining roadway, which belongs to the field of coupling support of mining roadways, and specifically comprises the following steps: acquiring rock mechanical parameters of a roadway to be detected, active support parameters of the roadway to be detected, passive support parameters of the roadway to be detected and field monitoring data; carrying out active and passive support strength coupling judgment on the front-section-exceeding roadway; determining that the safety coefficient of the active and passive combined support is within a reasonable range; determining the initial supporting force of the unit bracket; determining the deformation coupling of active and passive supports of the advanced roadway; determining that the reject ratio of the anchor rod is within a reasonable range; determining that the downward shrinkage amount of the unit support plunger does not exceed the limit; it is determined that the stable working posture of the unit support is not affected by side bulging; and the coupling support condition is observed, and finally whether the coupling support state is reached or not is determined. According to the method, a bimodal judgment mechanism of strength coupling and deformation coupling is established, a coupling regulation and control method is provided, and the strength stability and the deformation stability of the mining roadway are effectively guaranteed.

Description

Technical Field

[0001] The invention relates to the field of coupled support of mining tunnels, and specifically comprises a method for determining coupled support of an advanced mining tunnel. Background Art

[0002] The stability of the mining roadway is the key to safe and efficient mining in coal mines. With the advancement of deep coal mining, the surrounding rock stress on the roadway has increased significantly, especially in the advanced area of ​​the working face, which is affected by mining and has a highly concentrated stress, far exceeding the original rock stress level. In order to ensure the stability of the mining roadway, reasonable advanced support measures must be adopted to effectively disperse stress, control surrounding rock deformation, and improve the reliability and long-term bearing capacity of the support system.

[0003] The advanced support system of the mining tunnel is composed of two support forms: active support and passive support. Active support applies preload to support materials such as anchor rods and anchor cables to actively resist the surrounding rock load; passive support is that the unit support passively bears the external load by relying on its own strength and stiffness. Active-passive coupling support organically combines active support with passive support, comprehensively utilizes the high-strength stability of active support and the flexible adjustment ability of passive support, realizes the coupling of strength and deformation of the surrounding rock of the mining tunnel, active support system, and passive support system, limits the harmful deformation and damage of the tunnel and support structure, and achieves the purpose of integrated support, load uniformity, and stable mining tunnel.

[0004] Although the existing coupled support patents and technologies have certain advantages, they still have the following problems: ① Most of the research focuses on the research and development of new support materials and the innovation of support methods, and there are few studies on the determination method and implementation process of the coupling state between the active support system based on anchor support, the passive support system of unit support and the surrounding rock of the tunnel (such as the Chinese patent with publication number CN216767407U); ② The existing technology often improves the strength stability or deformation stability of the surrounding rock of the tunnel by coupling support alone, and does not consider the strength stability and deformation stability as two necessary conditions for achieving the overall stability of the tunnel (such as the Chinese patent with publication number CN114165269B); ③ Some technologies are only aimed at specific geological conditions or disaster phenomena (such as the Chinese patent with publication number CN222457475U), lack universality, and are difficult to be widely used in other complex mining environments. Based on this, a method for determining the coupled support of the advanced mining tunnel is urgently needed to solve the above problems. Summary of the invention

[0005] In view of the problems existing in the prior art, the present invention provides a method for determining coupled support of an advanced mining tunnel.

[0006] The present invention adopts the following technical solution: a method for determining coupled support of an advanced mining tunnel, comprising the following steps: S1: On-site investigation to obtain the rock mechanical parameters, active support parameters, passive support parameters and on-site monitoring data of the tunnel to be tested.

[0007] S2: Determination of coupling strength of active and passive support in the leading tunnel.

[0008] S2.1: Determine the safety factor of active and passive combined support Within a reasonable range.

[0009] S2.2: Determine the initial support force N of the unit support so that the passive support will not damage the tunnel roof.

[0010] S2.3: After the requirements of S2.1 and S2.2 are met, execute S3.

[0011] S3: Determination of coupling deformation of active and passive support in the leading tunnel.

[0012] S3.1: Determine that the anchor extension failure rate is within a reasonable range.

[0013] S3.2: Verify that the reduction in the unit support plunger does not exceed the limit.

[0014] S3.3: Ensure that bulging of the side panels does not affect the unit support’s ability to maintain a stable working posture.

[0015] S3.4: After the requirements of S3.1, S3.2, and S3.3 are met, perform S4.

[0016] S4: Conduct comprehensive observation on the coupled support situation and ultimately determine whether the coupled support state has been reached.

[0017] Preferably, the rock mechanical parameters of the tunnel to be tested in S1 include the rock properties and structure of the tunnel roof, and the rock properties and structure of the tunnel side; the active support parameters of the tunnel to be tested include the specification parameters of anchor rods and anchor cables, the layout spacing of anchor rods and anchor cables, and the layout quantity of anchor rods and anchor cables; the passive support parameters of the tunnel to be tested 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 include the deformation observation data of the tunnel surrounding rock, the extension length data of the anchor rods, and the downward shrinkage of the active column of the unit support.

[0018] Preferably, the active and passive combined support safety factor in S2.1 is for: .

[0019] Where: D is the spacing of the unit supports, m; B is the width of the tunnel section, m; d g is the anchor spacing, m; d sis the anchor cable spacing, m; n g is the number of single-row anchor rods, roots; n s is the number of single-row anchor cables, roots; F g Design anchoring force for a single anchor, kN; F s is the design anchoring force of a single anchor cable, kN; L is the effective anchoring length of the anchor, m; γ is the density of the rock formation anchored by the anchor, kN / m 3 ; F is the periodic breaking step of the basic top, m; K q is the load distribution factor, K q =D / 0.5F;m E is the thickness of the basic top, m; m Z is the thickness of the immediate roof, m; γ E is the average bulk density of the rock layer at the basic top, kN / m 3 ; γ Z is the average bulk density of the immediate top rock layer, kN / m 3 ; B0 is the bearing width of the tunnel, m; is the top breaking efficiency, ρ is the rock medium density, kg / m 3 ; C P is the 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 unit support top control area, Rated operating resistance for the unit bracket.

[0020] If K l ≥2, the safety factor of active and passive combined support Within a reasonable range.

[0021] Preferably, S2.2 specifically includes: Determine the initial support force N of the unit support so that the specific pressure of the tunnel roof is not greater than the compressive strength σ of the tunnel roof rock c .

[0022] The initial support force N of the unit support satisfies: .

[0023] Where: N is the initial support force of the unit support, kN; A is the effective contact area between the top beam and the top plate of the unit support, m 2 ; σ c is the compressive strength of the tunnel roof rock, MPa; k is the specific pressure safety factor.

[0024] Preferably, S3.1 specifically includes: the actual extension length ΔL of the anchor rod satisfies: .

[0025] If the anchor rod extension is good, the anchor rod extension is judged to be qualified; otherwise, the anchor rod extension is judged to be unqualified.

[0026] Where: y is the yield strength of the anchor material, MPa; E is the elastic modulus of the anchor material, MPa; L is the effective anchoring length of the anchor, m.

[0027] The anchor rod extension failure rate refers to the proportion of anchor rods that fail to meet the extension requirements to the total number of anchor rods. If the anchor rod extension failure rate is not greater than 3%, the anchor rod extension failure rate is within a reasonable range.

[0028] Preferably, S3.2 specifically includes: the limit shrinkage of the active column of the unit support Δh max for: Δh max =H max -H min .

[0029] Where: H max H is the maximum support height of the unit support, m; min is the minimum support height of the unit bracket, m.

[0030] The actual downward contraction Δh of the unit support plunger is equal to the contraction of the tunnel roof and floor plates.

[0031] Make sure that the actual shrinkage of the unit support plunger Δh does not exceed the limit shrinkage of the plunger Δh max Right now: Δh max＞ μΔh.

[0032] Where μ is the safety factor of shrinkage.

[0033] Preferably, S3.3 specifically includes: calculating the overturning moment of the unit support, where the overturning moment of the unit support is: .

[0034] M G is the self-weight overturning moment, 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 top plate of the unit support, and 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, and is the moment generated by the vertical force on the overturning edge of the unit support base; M f1 is the overturning moment of the friction force of the top plate of the unit support, and 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 overturning moment of the friction force of the unit support base, and is the moment generated by the friction force on the lower surface of the unit support base on the overturning edge; is the external overturning moment, which is the moment generated on the overturning edge by the horizontal force applied to the unit support by the bulging of the tunnel wall.

[0035] Determine that the overturning moment M of the unit support is greater than 200.

[0036] Preferably, S4 specifically includes: S4.1: Observe the deformation of the tunnel surrounding rock: After one week of implementation of the above-mentioned S1 to S3 coupling steps, observe the deformation of the tunnel surrounding rock for one week. Check the top, side and bottom surrounding rocks of the leading tunnel every day, and pay attention to whether there is obvious roof collapse, side bulging and bottom bulging. If the weekly cumulative deformation at the point where the tunnel surrounding rock has the largest deformation does not exceed 15mm, the deformation of the tunnel surrounding rock is considered normal.

[0037] S4.2: Observe the working condition of the active support system: One week after the implementation of the above-mentioned S1 to S3 coupling steps, observe the deformation of the active support for one week, check the status of the active support system every day, and pay attention to the elongation and anchoring condition of the anchor rods; if no more than two anchor rods have an extension length exceeding their maximum extension length within a week, and no anchor rod breakage occurs, the active support system is deemed to be working normally.

[0038] S4.3: Observe the working condition of the passive support system: One week after the above coupling steps S1 to S3 are implemented, observe the deformation of the passive support for one 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 a cumulative one week, the passive support system is considered to be working normally.

[0039] The present invention has the following beneficial effects: 1. The present invention provides a method for determining coupled support of advanced mining tunnels, which for the first time makes an overall coupled determination of active support, passive support and tunnel surrounding rock, and establishes a dual-mode determination mechanism of strength coupling and deformation coupling. This mechanism breaks through the limitation of traditional determination methods that only focus on strength or deformation, and ensures the scientificity and comprehensiveness of the determination results.

[0040] 2. The method for determining the coupled support of the advanced mining tunnel provided by the present invention uses the coefficient correction and theoretical calculation methods based on field data to provide a scientific and quantitative coupling determination basis. Compared with the traditional empirical judgment method, this method has stronger operability and effectively improves the safety of the overall support of the mining tunnel.

[0041] 3. The method for determining the coupled support of the advanced mining tunnel provided by the present invention monitors the deformation of the anchor group in the tunnel to be tested in real time through the optical fiber synchronous deformation composite anchor. The device has a fully distributed and high-precision monitoring capability, can adapt to the complex underground environment and realize long-term stable monitoring; a unit support overturning instability assessment method is also established to achieve quantitative assessment of the instability state of the unit support, which can guide the scientific use of the unit support.

[0042] 4. The coupling determination method provided by the present invention realizes the closed-loop management of "evaluation-warning-control", and accurately formulates differentiated control strategies by accurately evaluating the coupling state of the support system. 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 mine safety, production efficiency and economic benefits. BRIEF DESCRIPTION OF THE DRAWINGS

[0043] Figure 1 Schematic diagram of the data acquisition system for the actual extension length of the anchor.

[0044] Figure 2 Schematic diagram of the optical fiber synchronous deformation composite anchor.

[0045] Figure 3 It is a flow chart of the present invention.

[0046] Figure 4 Schematic diagram of unit support instability and overturning.

[0047] Figure 5 This is the force analysis diagram when the unit support becomes unstable and overturns.

[0048] Figure 6 This is the force analysis diagram of the unit bracket in Example 1. DETAILED DESCRIPTION

[0049] The specific implementation of the present invention is further described below in conjunction with the accompanying drawings and specific embodiments: Figures 1 to 6 A method for determining coupled support of an advanced mining tunnel comprises the following steps: S1: On-site investigation to obtain the rock mechanical parameters, active support parameters, passive support parameters and on-site monitoring data of the tunnel to be tested.

[0050] The rock mechanical parameters of the tunnel to be tested include the rock properties and structure of the tunnel roof, and the rock properties and structure of the tunnel side. The active support parameters of the tunnel to be tested include the specification parameters of anchor rods and anchor cables, the layout spacing of anchor rods and anchor cables, and the layout quantity of anchor rods and anchor cables. The passive support parameters of the tunnel to be tested 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 include the deformation observation data of the tunnel surrounding rock, the extension length data of the anchor rods, and the shrinkage of the active column of the unit support.

[0051] S2: Determination of coupling strength of active and passive support in the leading tunnel.

[0052] S2.1: Determine the safety factor of active and passive combined support Within a reasonable range.

[0053] Calculate the maximum support strength that the anchor active support can provide, the maximum passive support strength that the unit support can provide, and the tunnel roof load, and determine whether the maximum support strength that the active and passive combined support can provide can resist the maximum roof load, and then determine the active and passive combined safety factor K l Within a reasonable range.

[0054] Active and passive combined safety factor K l The calculation principle is: .

[0055] Then we get: .

[0056] Where: p l The maximum support strength that the active and passive combined support system can provide, MPa; S k is the top 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 load, kN; Q3 is the impact load, kN.

[0057] Final active and passive combined support safety factor for: ; Where: D is the spacing of the unit supports, m; B is the width of the tunnel section, m; d g is the anchor spacing, m; d s is the anchor cable spacing, m; n g is the number of single-row anchor rods, roots; n s is the number of single-row anchor cables, roots; F g Design anchoring force for a single anchor, kN; F s is the design anchoring force of a single anchor cable, kN; L is the effective anchoring length of the anchor, m; γ is the density of the rock formation anchored by the anchor, kN / m 3 .

[0058] F is the periodic breaking step of the basic top, m; K q is the load distribution factor, K q =D / 0.5F;m E is the thickness of the basic top, m; m Z is the thickness of the immediate roof, m; γ E is the average bulk density of the rock layer at the basic top, kN / m 3 ; γ Z is the average bulk density of the immediate top rock layer, kN / m 3 ; B0 is the bearing width of the roadway, m; is the top breaking efficiency, ρ is the rock medium density, kg / m 3 ; C P is the 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 unit support top control area, Rated operating resistance for the unit bracket.

[0059] Safety factor K of active and passive combined support l The judgment, evaluation and treatment measures are shown in Table 1.

[0060] Table 1

[0061] Reinforced support: ① Increase the density of anchor support, that is, add more anchors; ② Increase the strength of passive support, that is, reduce the spacing between unit brackets.

[0062] If K l ≥2, the safety factor of active and passive combined support Reasonable. If the safety factor of active and passive combined support If it is too small, the safety factor of active and passive combined support should be increased by reinforcing the support. Adjust within a reasonable range.

[0063] S2.2: Determine the initial support force N of the unit support so that the passive support will not damage the tunnel roof.

[0064] Setting too large an initial support force N at the initial support stage of the unit support will cause the top beam of the unit support to damage the rock layer of the tunnel roof, resulting in the phenomenon of "drilling the top" of the unit support. In order to prevent the occurrence of the drilling top phenomenon, it is necessary to determine a reasonable initial support force N of the unit support so that the specific pressure of the tunnel roof is not greater than the compressive strength of the rock on the tunnel roof σ c .

[0065] The specific pressure of the tunnel roof is the ratio of the initial support force of the unit support to the area of ​​the top beam of the unit support. To ensure that the specific pressure of the tunnel roof is not greater than the compressive strength of the roof rock σc, the initial support force N of the unit support must satisfy: .

[0066] Where: N is the initial support force of the unit support, kN; A is the top beam area of ​​the unit support, m 2 ; σ c is the compressive strength of the tunnel roof rock, MPa; k is the specific pressure safety factor, usually 2~3.

[0067] If N≥(Aσ c ) / k, the initial support force of the unit support should be reduced until N < (Aσ c) / k. The premise of reducing the initial support force of the unit support is to strictly follow the coal mine management regulations and meet the initial support force requirements.

[0068] S2.3: After the requirements of S2.1 and S2.2 are met, execute S3.

[0069] That is, the active and passive support strength coupling judgment of the leading tunnel needs to meet K l ≥2, .

[0070] S3: Determination of coupling deformation of active and passive support in the leading tunnel.

[0071] S3.1: Determine whether the anchor elongation is within reasonable limits.

[0072] The maximum extension length of the anchor refers to the maximum axial deformation that the anchor can produce before it enters the plastic stage. The deformation of the anchor entering the plastic stage means that the deformation of the anchor has exceeded the elastic limit, and the anchor no longer fully recovers its original shape. Even after the load is removed, it will leave some permanent deformation. This indicates that the bearing capacity of the anchor has decreased, or the anchor is close to its failure limit. For the support system, the anchor entering the plastic stage may affect the stability of the support, especially when the anchor continues to bear excessive loads, which may lead to support failure or further deformation of the surrounding rock mass.

[0073] Therefore, when designing the support system, it is necessary to ensure that the deformation of the anchor does not enter the plastic stage, or to take appropriate measures when necessary to ensure that the support system can still provide effective support to prevent rock collapse or support failure.

[0074] Determine the reasonable range of the actual extension length ΔL of the anchor rod, that is, the actual extension length ΔL of the anchor rod satisfies: .

[0075] Where: y is the yield strength of the anchor material, MPa; E is the elastic modulus of the anchor material, MPa; L is the effective anchoring length of the anchor, m.

[0076] If the above conditions are met, the anchor extension is judged to be qualified, otherwise the anchor extension is judged to be unqualified.

[0077] It should be noted that in the mining roadway, the anchor rod does not exist alone, but the group of anchor rods plays a supporting role together. The damage or failure of a single anchor rod usually does not cause the overall support failure of the roadway or the loss of control of the surrounding rock. Therefore, the object of S3.1 monitoring is the group of anchor rods in the mining roadway. The anchor rod extension failure rate refers to the proportion of the number of anchor rods with unqualified anchor rod extension to the total number of anchor rods. If the anchor rod extension failure rate is not greater than 3%, the anchor rod extension failure rate is within a reasonable range.

[0078] If the anchor rod extension failure rate exceeds 3%, an additional anchor rod shall be driven next to the anchor rod with unqualified extension length.

[0079] There are a large number of anchor rods in the mining tunnel. If the actual extension length ΔL of each anchor rod is monitored, the workload is huge and the monitoring cost is extremely high. In order to solve the above problems, the present invention designs an optical fiber synchronous deformation composite anchor rod, such as Figure 1 and 2 , the optical fiber synchronous deformation composite anchor is used to monitor the actual extension length ΔL of the anchor group in real time.

[0080] Production and assembly steps of optical fiber synchronous deformation composite anchor: Step 1: Anchor rod grooving: along the axial direction of anchor rod 1 and parallel to the center axis of the anchor rod, micro groove 2 is opened on the surface of the anchor rod. The depth of micro groove 2 is ≤3% of the anchor rod diameter and the width of micro groove 2 is 1-1.5mm.

[0081] Step 2: Bury the optical fiber: Select a Brillouin distributed optical fiber sensor 3 with an optical fiber diameter of 100~250μm; lay the optical fiber in the microgroove 2 to ensure that the optical fiber is taut and without bends; reserve a 5cm optical fiber buffer at the anchoring end of the anchor rod, and use a protective cover 4 to seal and protect the optical fiber end; reserve an installation position for the optical fiber connector 8 at the exposed end of the anchor rod to connect it to the data acquisition system.

[0082] Step 3: Backfill the microgroove with high-strength filling material: Use carbon fiber reinforced epoxy resin material 5 to fill the microgroove to ensure that the optical fiber fits tightly to the anchor rod; after the filling material is cured, grind the surface of the anchor rod to ensure that the surface of the anchor rod is flat to avoid the protrusion of the filling area affecting the anchoring of the anchor rod to the surrounding rock.

[0083] Step 4: Anchor installation: Use an anchor drill to drill holes at the marked hole positions on the tunnel roof 12; after drilling, insert one fast anchor and one medium-speed anchor in sequence; connect the anchor to the anchor mixer, insert the anchor into the hole and slowly stir the anchor for 25-35 seconds and 90 seconds for gelling; after waiting for 8 minutes, tighten the nut 7 with the anchor drill to press the anchor tray 6 against the rock surface; use a jackhammer or torque wrench to tighten the anchor for a second time to achieve the designed preload torque.

[0084] The data acquisition system for the actual extension length of the anchor rod is as follows: Figure 1 , its assembly order: Optical fiber synchronous deformation composite anchor → optical fiber end reserved at the exposed end of the anchor → optical fiber connector 8 → Brillouin optical time domain reflectometer equipment 9 → data acquisition card 10 → computer 11.

[0085] How to use the anchor bolt actual extension length data acquisition system: ① After all the equipment is turned on, the Brillouin optical time domain reflectometer 9 scatters light source to the Brillouin distributed optical fiber sensor in the anchor rod.

[0086] ② The Brillouin distributed optical fiber sensor 3 measures the change of the Brillouin frequency shift and calculates the strain value of the optical fiber.

[0087] ③ The Brillouin optical time domain reflectometer 9 receives and processes the signal returned by the optical fiber sensor.

[0088] 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.

[0089] ⑤ Analyze the actual deformation of the anchor rod through computer software 11, and output and display the actual deformation result of the anchor rod.

[0090] S3.2: Verify that the reduction in the unit support plunger does not exceed the limit.

[0091] As the mine pressure appears, the tunnel cross-section converges and the top and bottom plates move closer. If the tunnel convergence exceeds the limit of the unit support live column shrinkage, the unit support will be "crushed to death", causing damage to the support equipment and seriously endangering the safety of life and property underground. The necessary condition to avoid the unit support from being crushed is that the actual shrinkage of the unit support live column Δh cannot exceed the limit shrinkage of the live column Δh max , the calculation process is as follows: The limit shrinkage of the unit support column Δh max for: Δh max =H max -H min .

[0092] Where: H max H is the maximum support height of the unit support, m; min is the minimum support height of the unit bracket, m.

[0093] The actual downward shrinkage amount Δh of the unit support plunger is determined by on-site monitoring, and the actual downward shrinkage amount Δh of the unit support plunger is equal to the shrinkage amount of the top and bottom plates of the tunnel.

[0094] Make sure that the actual shrinkage of the unit support plunger Δh does not exceed the limit shrinkage of the plunger Δh max Right now: Δh max＞ μΔh.

[0095] In the formula, μ is the safety factor of shrinkage, which is taken as 1.2-1.5.

[0096] If Δh max≤ μΔh, then the support density of the unit support is strengthened so that Δh max＞ μΔh, strengthening the support density of the unit bracket is to increase the number of unit brackets and reduce the spacing between unit brackets.

[0097] S3.3: Ensure that bulging of the side panels does not affect the unit support’s ability to maintain a stable working posture.

[0098] As the mining face advances, the stress of the tunnel surrounding rock is redistributed, and the surrounding rock stress on the side increases. Due to insufficient support strength and other reasons, the surrounding rock of the side gradually bulges out and squeezes the unit support, which is prone to gradual instability of the unit support, and eventually causes the unit support to overturn and become unstable in the tunnel. The fundamental reason for the overturning and instability of the unit support is that the surrounding rock of the side gradually bulges out and squeezes the unit support, and the horizontal force applied to the unit support by the side destroys the balanced working state of the unit support.

[0099] The risk of unit support toppling and instability can be judged by the unit support overturning moment M. The unit support overturning moment refers to the product of the overturning load on the unit support and the overturning arm (the distance from the overturning load to the overturning edge). According to the different overturning loads, the unit support overturning moment is divided into the unit support self-weight overturning moment, the unit support support overturning moment, and the unit support external force overturning moment. The unit support self-weight overturning moment refers to the overturning moment M generated by the unit support's own gravity G on the overturning edge. G The overturning moment of the unit support refers to the moment M generated on the overturning edge 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 after the unit support applies support resistance to the top and bottom plates of the tunnel. N1 、M N2 、M f1 、M f2 The external overturning moment refers to the horizontal force F applied to the unit support by the bulging of the tunnel side. X The overturning moment M generated on the overturning edge Fx The combined moment refers to the vector sum of the above moments. The unit support is kept stable by the combined moment. The key to preventing the unit support from tipping over is to ensure that the combined tipping moment is greater than 0 (clockwise moment is defined as positive).

[0100] Calculate the combined overturning moment of the unit support. The combined overturning moment of the unit support is: .

[0101] The overturning arm refers to the distance from the overturning load to the overturning edge. Figure 5 As shown. G Half of the width of the unit support base plate, m; H N1 It is half of the width of the top beam of the unit support, m; N2 H is half of the width of the unit support base, m; f1 H is the unit support height (tunnel height), m; f2 0m; H FxIt is half of the unit support height (tunnel height), m.

[0102] M G is the overturning moment generated by the deadweight G of the unit support on the overturning edge, M N1 M is the moment generated by the vertical force on the overturning edge of the unit support top beam, N2 M is the moment generated by the vertical force on the overturning edge of the unit support base. f1 M is the moment generated by the friction force on the upper surface of the top beam of the unit support on the overturning edge. f2 is the moment generated by the friction force on the lower surface of the unit support base on the overturning edge; It is the moment on the overturning edge caused by the horizontal force applied to the unit support by the bulging of the tunnel wall.

[0103] Determine whether the unit bracket overturning moment M is greater than 200.

[0104] If M is less than or equal to 200, the unit support is lowered to adjust its posture, and the surrounding rock of the side is trimmed so that it no longer protrudes and squeezes the unit support.

[0105] S3.4: After the requirements of S3.1, S3.2, and S3.3 are met, perform S4.

[0106] S4: Conduct comprehensive observation on the coupled support situation and ultimately determine whether the coupled support state has been reached.

[0107] S4.1: Observe the deformation of the tunnel surrounding rock: After one week of implementation of the above-mentioned S1 to S3 coupling steps, observe the deformation of the tunnel surrounding rock for one week. Check the top, side and bottom surrounding rocks of the leading tunnel every day, and pay attention to whether there is obvious roof collapse, side bulging and bottom bulging. If the weekly cumulative deformation at the point where the tunnel surrounding rock has the largest deformation does not exceed 15mm, the deformation of the tunnel surrounding rock is considered normal.

[0108] S4.2: Observe the working condition of the active support system: One week after the implementation of the above-mentioned S1 to S3 coupling steps, observe the deformation of the active support for one week, check the status of the active support system every day, and pay attention to the elongation and anchoring condition of the anchor rods; if no more than two anchor rods have an extension length exceeding their maximum extension length within a week, and no anchor rod breakage occurs, the active support system is deemed to be working normally.

[0109] S4.3: Observe the working condition of the passive support system: One week after the above coupling steps S1 to S3 are implemented, observe the deformation of the passive support for one 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 a cumulative one week, the passive support system is considered to be working normally.

[0110] If the above observations are normal, it is determined that the coupled support state has been reached.

[0111] Example 1: The following is a verification and explanation of the above-mentioned method for determining the coupling of the leading support of the mining roadway through a specific construction site case of the 2309 working face track drift of a coal mine: S1: On-site investigation to obtain the rock mechanical parameters, active support parameters, passive support parameters and on-site monitoring data of the tunnel to be tested.

[0112] Basic parameters and pressure of tunnel rock formations:

[0113] The track tunnel is driven along the goaf, and the coal pillar width is B m =4m, arranged along the coal seam floor supporting the top coal; rectangular section, the net width of the section is B, 5.6m, and the net height is 3.8m. The elastic-plastic boundary length of the solid coal seam is B s =3.15m; Solid coal density γ: 14kN / m 3 ; Direct top thickness m Z =4.48m, average bulk density of direct top γ Z =23kN / m 3 ; Thickness of basic top m E =11.51m; average bulk density of basic top γ E =23kN / m 3 .

[0114] The periodic breaking step of the basic top is F=13m; the medium density ρ=23kg / m 3 , P wave velocity C P =5.8×10 3 m / s, vibration speed v´ P =1.37m / s.

[0115] Active support situation: Anchor type and specification: Φ22×2800mm left-handed threaded steel anchor without longitudinal ribs, single anchor design anchoring force F g =190kN, yield strength of anchor material σ y =500MPa; moment of inertia of anchor section I=11493.185mm 4 .

[0116] Anchor arrangement: Anchor spacing is 0.9m, row spacing d g =0.9m, number of anchor rods in a single row n g =7, L is the effective anchoring length of the anchor rod 2.65m; anchor rod anchoring rock density γ = 13kN / m 3 .

[0117] Anchor cable type and specifications: Φ21.8mm×8200mm 1×19S structural steel strand anchor cable. Number of single-row anchor cables n s = 3 anchors, single anchor cable design anchoring force F s =553kN.

[0118] Anchor cable arrangement: spacing 1.8m, row spacing d s =1.8m.

[0119] Passive support: ZQ4000 / 20.6 / 45 type unit support is used to support the top plate, with a rated working resistance of N d =4000kN; minimum support height H min =2.06m; Maximum support height H max =4.5m; top beam size: 1600mm×780mm, base size: 1812mm×940mm. The effective contact area between the top beam and the top plate of the unit support is A=1.248m 2 ; Double row arrangement of unit brackets n d =2 frames, the spacing between unit supports is D=6m, and the initial support force is set to 600kN during support.

[0120] S2: Determination of coupling strength of active and passive support in the leading tunnel.

[0121] S2.1: Safety factor K of active and passive combined support l calculate: =2.4.

[0122] 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.

[0123] K l =2.4>2,K l Within a reasonable range.

[0124] S2.2: Determine the initial support force N of the unit support so that the passive support will not damage the tunnel roof.

[0125] =6364.8kN.

[0126] 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.

[0127] 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.

[0128] 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.

[0129] S3: Determination of coupling deformation of active and passive support in the leading tunnel.

[0130] S3.1: Determine that the anchor extension failure rate is within a reasonable range.

[0131] Determine the reasonable range of the actual extension length ΔL of the anchor rod, that is: =1.325cm.

[0132] 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.

[0133] 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.

[0134] S3.2: Verify that the reduction in the unit support plunger does not exceed the limit.

[0135] The limit shrinkage of the unit support column Δh max =H max -H min =2.44m.

[0136] On-site monitoring data show that the maximum shrinkage of the top and bottom plates of the track trough of the 2309 working face is 0.2m, that is, the actual shrinkage of the unit support column is Δh=0.2m. Obviously: Δh max＞ μΔh=0.3m.

[0137] Note: In the formula, μ is the safety factor of shrinkage, which is taken as 1.5.

[0138] The actual height of the tunnel section must be greater than the minimum height of the unit support. The unit support will not be crushed in the 2309 track drift and subsequent monitoring will be normal.

[0139] S3.3: Ensure that bulging of the side panels does not affect the unit support’s ability to maintain a stable working posture.

[0140] The steady-state analysis method of the unit bracket is as follows: In order to calculate the overturning moment of the unit support, the unit support is taken as the research object for force analysis. Figure 6 :

[0141] =4003.38>200.

[0142] Where: self-weight overturning moment M G : refers to the overturning moment generated by the unit support's own gravity G on the overturning edge, where gravity G=54kN, overturning arm H G =0.47m.

[0143] Unit support top plate self-supporting overturning moment M N1 : It refers to the moment of force generated by the vertical force on the overturning edge of the unit support beam after the unit support applies support resistance to the top and bottom plates of the tunnel, where N1=4000kN and the overturning arm H N1 =0.47m.

[0144] Unit support base self-supporting overturning moment M N2 : It refers to the moment of force generated by the vertical force on the overturning edge of the unit support base after the unit support applies support resistance to the top and bottom plates of the tunnel, where N2=4000kN and the overturning arm H N2 =0.47m.

[0145] Unit support top plate friction overturning moment M f1 : It refers to the moment of force generated by the friction force on the upper surface of the top beam of the unit support on the overturning edge after the unit support applies support resistance to the top and bottom plates of the tunnel, where f1=μN1=1200kN, and the overturning arm H f1 =3.6m.

[0146] Unit bracket base friction overturning moment M f2 : It refers to the moment of force generated by the friction force on the lower surface of the unit support base on the overturning edge after the unit support applies support resistance to the top and bottom plates of the tunnel, where f2=μN2=1200kN, and the overturning arm H f1 =0m.

[0147] External overturning moment M Fx : refers to the moment of force exerted on the overturning edge by the horizontal force applied to the unit support by the bulging of the roadway side, where F X =190kN, overturning arm H Fx =1.8m.

[0148] ① The bending moment is calculated in a clockwise direction; the unit support loses stability and overturns into the tunnel. The overturning edge is the side of the unit support base in the tunnel, and the overturning point is the lowest point of the side of the unit support base in the tunnel.

[0149] ②Unit support weight 5400kg, gravity G=54kN, gravity acceleration g is 10m / s 2; The top beam and base of the unit support are subjected to the support reaction force of the surrounding rock. Considering the most dangerous situation, the unit support outputs a working resistance of 4000kN, that is, N1=N2=4000kN; there is a friction force f1 between the top beam of the unit support and the top plate rock, f1=μN1=1200kN; there is a friction force f2 between the base of the unit support and the bottom plate rock, f2=μN2=2000kN, μ is the friction coefficient and is taken as 0.5; the horizontal force F applied to the unit support by the bulging of the tunnel side x =190kN.

[0150] The overturning moment M of the unit support should be greater than 200, judging that the unit support of the 2309 working face track lane can maintain a steady state and be in a safe state, and subsequent normal monitoring is required.

[0151] In summary, the track lane of the 2309 working face meets the deformation coupling standard.

[0152] S4: Conduct comprehensive observation on the coupled support situation and ultimately determine whether the coupled support state has been reached.

[0153] After the implementation of the above S1-S3 coupling steps for one week, the deformation of the tunnel surrounding rock, the working condition of the active support system, and the working condition of the passive support system were observed for one week in the 2309 working face track tunnel. The observation results showed that: ① The cumulative deformation of the tunnel surrounding rock was 8mm at the maximum deformation point; ② There was no anchor rod whose extension length exceeded its maximum extension length, and no anchor rod fracture occurred; ③ There was no crushing of the support and instability and overturning of the unit support. Finally, it was determined that the tunnel reached the coupled support state.

[0154] 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 technicians in this technical field within the essential scope of the present invention should also fall within the protection scope of the present invention.

Claims

1. A method for determining coupled support of advanced mining tunnel, characterized in that: The following steps are involved: S1: On-site investigation to obtain rock mechanical parameters, active support parameters, passive support parameters and on-site monitoring data of the tunnel to be tested; S2: Determination of coupling strength of active and passive support in the advanced tunnel; S2.1: Determine the safety factor of active and passive combined support within reasonable limits; S2.2: Determine the initial support force N of the unit support so that the passive support will not damage the tunnel roof; S2.3: After meeting the requirements of S2.1 and S2.2, execute S3; S3: Determination of coupling deformation of active and passive support in the leading tunnel; S3.1: Determine that the anchor extension failure rate is within a reasonable range; S3.2: Ensure that the reduction of the unit support column does not exceed the limit; S3.3: Ensure that the bulging of the side parts does not affect the unit support 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 on the coupled support situation and ultimately determine whether the coupled support state has been reached.

2. The method for determining coupled support of advanced mining tunnel according to claim 1 is characterized in that: The rock mechanical parameters of the tunnel to be tested in S1 include the rock properties and structure of the tunnel roof, the rock properties and structure of the tunnel side, the active support parameters of the tunnel to be tested include the specification parameters of anchor rods and anchor cables, the layout spacing of anchor rods and anchor cables, and the layout quantity of anchor rods and anchor cables, the passive support parameters of the tunnel to be tested 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 include the deformation observation data of the tunnel surrounding rock, the extension length data of the anchor rods, and the shrinkage of the active column of the unit support.

3. The method for determining coupled support of advanced mining tunnel according to claim 1 is characterized in that: Safety factor of active and passive combined support in S2.1 for: ； Where: D is the spacing of the unit supports, m; B is the width of the tunnel section, m; d g is the anchor spacing, m; d s is the anchor cable spacing, m; n g is the number of single-row anchor rods, roots; n s is the number of single-row anchor cables, roots; F g Design anchoring force for a single anchor, kN; F s is the design anchoring force of a single anchor cable, kN; L is the effective anchoring length of the anchor, m; γ is the density of the rock formation anchored by the anchor, kN / m 3 ; F is the periodic breaking step of the basic top, m; K q is the load distribution factor, K q =D / 0.5F;m E is the thickness of the basic top, m; m Z is the thickness of the immediate roof, m; γ E is the average bulk density of the rock layer at the basic top, kN / m 3 ; γ Z is the average bulk density of the immediate top rock layer, kN / m 3 ; B0 is the bearing width of the tunnel, m; is the top breaking efficiency, ρ is the rock medium density, kg / m 3 ; C P is the 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 unit support top control area, Rated working resistance of the unit support; If K l ≥2, the safety factor of active and passive combined support Within a reasonable range.

4. The method for determining coupled support of advanced mining tunnel according to claim 1 is characterized in that: S2.2 specifically includes: Determine the initial support force N of the unit support so that the specific pressure of the tunnel roof is not greater than the compressive strength σ of the tunnel 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 and the top plate of the unit support, m 2 ; σ c is the compressive strength of the tunnel roof rock, MPa; k is the specific pressure safety factor.

5. The method for determining coupled support of advanced mining tunnel according to claim 1, characterized in that: S3.1 specifically includes: The actual extension length ΔL of the anchor bolt satisfies: ； The anchor extension is judged to be qualified, otherwise the anchor extension is judged to be unqualified; Where: y is the yield strength of the anchor material, MPa; E is the elastic modulus of the anchor material, MPa; L is the effective anchoring length of the anchor, m; The anchor rod extension failure rate refers to the proportion of anchor rods that fail to meet the extension requirements to the total number of anchor rods. If the anchor rod extension failure rate is not greater than 3%, the anchor rod extension failure rate is within a reasonable range.

6. The method for determining coupled support of advanced mining tunnel according to claim 1, characterized in that: S3.2 specifically includes: The limit shrinkage of the unit support column Δh max for: Δh max =H max -H min ; Where: H max H is the maximum support height of the unit support, m; min is the minimum support height of the unit bracket, m; The actual downward shrinkage of the unit support column Δh is equal to the shrinkage of the top and bottom plates of the tunnel; Make sure that the actual shrinkage of the unit support plunger Δh does not exceed the limit shrinkage of the plunger Δh max Right now: Δh max＞ μΔh; Where μ is the safety factor of shrinkage.

7. The method for determining coupled support of advanced mining tunnel according to claim 1, characterized in that: S3.3 specifically includes: calculating the overturning moment of the unit support, the overturning moment of the unit support is: ； M G is the self-weight overturning moment, 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 top plate of the unit support, and 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, and is the moment generated by the vertical force on the overturning edge of the unit support base; M f1 is the overturning moment of the friction force of the top plate of the unit support, and 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 overturning moment of the friction force of the unit support base, and is the moment generated by the friction force on the lower surface of the unit support base on the overturning edge; is the external overturning moment, which is the moment generated by the horizontal force applied to the unit support by the bulging of the tunnel side on the overturning edge; Determine that the overturning moment M of the unit support is greater than 200.

8. The method for determining coupled support of advanced mining tunnel according to claim 1 is characterized in that: S4 specifically includes: S4.1: Observe the deformation of the tunnel surrounding rock: After the above S1 to S3 coupling steps are implemented for one week, observe the deformation of the tunnel surrounding rock for one week. Check the top, side and bottom surrounding rocks of the leading tunnel every day, and pay attention to whether there is obvious roof collapse, side bulging and bottom bulging. If the accumulated deformation of the tunnel surrounding rock does not exceed 15mm at the place with the largest deformation, the deformation of the tunnel surrounding rock is considered normal; S4.2: Observe the working condition of the active support system: After the above S1 to S3 coupling steps are implemented for one week, observe the deformation of the active support for one week, check the status of the active support system every day, and pay attention to the elongation and anchoring of the anchor rods; if no more than two anchor rods have an extension length exceeding their maximum extension length within a week, and no anchor rod fracture occurs, the active support system is considered to be working normally; S4.3: Observe the working condition of the passive support system: One week after the above coupling steps S1 to S3 are implemented, observe the deformation of the passive support for one 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 a cumulative one week, the passive support system is considered to be working normally.

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