A roadway roof static cracking method based on while-drilling detection

Abstract

The application discloses a roadway roof static cracking method based on while-drilling detection, which comprises the following steps: drilling a hole in the roadway roof, and collecting while-drilling parameters in the drilling process and longitudinal wave velocity and transverse wave velocity in the hole in real time; determining rock tensile strength corresponding to different depth ranges of the hole based on the while-drilling parameters, the longitudinal wave velocity and the transverse wave velocity; according to the different depth ranges and the corresponding rock tensile strength, solid static cracking cartridges matching the length of each depth range are prepared, and the water-cement ratio and early strength agent dosage of the corresponding cartridges are adjusted according to the rock tensile strength of each depth range; and the adjusted segmented cartridges are loaded into the corresponding depth ranges of the hole and the hole is sealed; thereby, the while-drilling parameters, the longitudinal wave velocity and the transverse wave velocity are integrated, and a constructed rock uniaxial compressive strength inversion model and a rock tensile strength analysis model are used for calculation, so that the effect of in-situ and rapid acquisition of rock tensile strength at a construction site underground is achieved.

Description

Technical Field

[0001] This invention relates to the field of underground engineering technology, and in particular to a method for statically fracturing the roof of a tunnel based on drilling detection. Background Technology

[0002] In mining, especially coal mining, when there are hard rock strata in the roof of roadways, they often cannot collapse naturally and in a timely manner as the working face advances. This "suspended roof" phenomenon is a major cause of abnormally severe mine pressure in mining roadways, increased risk of spontaneous combustion of coal left in goaf areas, and easy accumulation of excessive gas in the upper corner of the working face. In order to ensure safe production in the mine, it is necessary to artificially fracture the hard overlying rock structure of the roadway roof to promote its timely collapse.

[0003] Currently, commonly used fracturing techniques for hard overburden in underground mine roadways mainly include explosive blasting, high-pressure water injection weakening, and carbon dioxide phase change fracturing. These techniques have achieved certain results under specific conditions, but they also have obvious limitations. For example, explosive blasting produces strong vibrations, shock waves, and open flames, posing a significant safety hazard in high-gas mines; high-pressure water injection has high requirements for water supply and rock permeability, making it difficult to implement in arid or dense rock formations; and carbon dioxide phase change fracturing is costly and requires extremely strict borehole sealing.

[0004] Therefore, static fracturing technology, as a novel method for inducing rock fracturing, relies on the expansion stress generated by the hydration reaction of static fracturing agents (usually calcium oxide-based materials) to slowly fracture the rock. It boasts significant advantages such as a stable process, no impact or vibration, no open flame, and no harmful gas generation, and has been gradually applied in underground rock strata control in recent years. However, in the current application of static fracturing technology for fracturing roadway roofs, key process parameters, such as the water-cement ratio of the static fracturing agent and the amount of early-strength agents (such as calcium chloride and sodium sulfate), are mostly set based on the experience of on-site engineers, lacking scientific and quantitative design basis.

[0005] The fundamental reason lies in the inability to quickly, accurately, and in situ obtain key mechanical parameters of the roof strata—especially the tensile strength of the rock—at the construction site. Rock tensile strength is a direct mechanical criterion determining whether the expansion stress of the static fracturing agent can be successfully overcome, leading to tensile failure of the rock. Currently, traditional methods for obtaining rock tensile strength require core drilling, transporting the cores to a surface laboratory for tests such as Brazilian fracturing. This process is time-consuming and costly, and disturbances during core drilling reduce the representativeness of the test results, completely failing to meet the real-time and economical requirements of downhole engineering.

[0006] Therefore, the lack of a method for rapidly, in-situ, and accurately obtaining the tensile strength of the roadway roof rock in underground operations has become a core bottleneck restricting the intelligent and precise development of static fracturing technology. To solve this problem, there is an urgent need to propose an efficient static fracturing method. Summary of the Invention

[0007] The technical problem solved by this invention is: how to obtain the tensile strength of the roof rock along the drilling depth direction in real time, in situ and continuously at the site of underground tunnel engineering, avoiding reliance on time-consuming and labor-intensive core sampling laboratory tests.

[0008] To solve the above technical problems, the present invention provides the following technical solution: a method for static fracturing of roadway roof based on drilling detection, comprising the following steps: S1, drilling a hole into the roadway roof and collecting drilling parameters and longitudinal wave velocity and transverse wave velocity in the hole in real time during the drilling process, wherein the drilling parameters include drilling rig thrust, drill bit rotation speed, drilling torque, drilling speed and drilling depth.

[0009] S2. Based on the drilling parameters, the longitudinal wave velocity, and the transverse wave velocity, determine the rock tensile strength corresponding to different depth ranges of the borehole.

[0010] S3. Based on the different depth ranges and their corresponding rock tensile strengths, prepare solid static fracturing cartridges that match the length of each depth range, and adjust the water-cement ratio and early strength agent dosage of the corresponding cartridges according to the rock tensile strength of each depth range to control the expansion stress and hydration expansion time of each segment of the cartridge.

[0011] S4. Load the adjusted segmented cartridges into the depth range corresponding to the drill hole and seal the hole.

[0012] Preferably, step S2 includes: S21, standardizing and reducing the dimensions of the collected drilling parameters, P-wave velocity and S-wave velocity to obtain a standard dataset.

[0013] S22. Based on the standard dataset and the pre-built rock uniaxial compressive strength inversion model, determine the rock uniaxial compressive strength along the depth direction of the borehole.

[0014] S23. Based on the longitudinal wave velocity and the transverse wave velocity, calculate the dynamic bulk modulus and dynamic Poisson's ratio of the rock.

[0015] S24. Based on the uniaxial compressive strength of the rock, the dynamic bulk modulus, and the dynamic Poisson's ratio, determine the tensile strength of the rock using a pre-constructed rock tensile strength analysis model.

[0016] Preferably, in step S21, the t-SNE algorithm is used to reduce the dimensionality of the data.

[0017] Preferably, in step S22, the calculation formula for the uniaxial compressive strength inversion model of the rock is as follows: ;in, It represents the uniaxial compressive strength of the rock. For drilling rig propulsion, The linear attenuation coefficient of tensile stress is... The angle between the drill bit and the rock; The width of the drill bit's cutting edge. The rock cutting coefficient, The circumference of the drill bit. This is an empirical coefficient, usually taken as 0.1; For drilling speed, This represents the drill bit rotation speed.

[0018] Preferably, in step S24, the dynamic bulk modulus The calculation formula is as follows: ;in, The density of the rock can be obtained by core sampling or assigned a value based on lithological experience. For longitudinal wave velocity, This refers to the transverse wave velocity.

[0019] The dynamic Poisson ratio The calculation formula is as follows: .

[0020] The calculation formula for the rock tensile strength analysis model is as follows: ;in, , and The fitting coefficients were obtained through fitting analysis of a large amount of indoor experimental data.

[0021] Preferably, in step S2, the strategy for dividing the different depth ranges is as follows: based on the curve of the change of the tensile strength of the rock with the drilling depth, the continuous segment with the curve amplitude fluctuation of less than 3% is divided into the same depth range.

[0022] Preferably, in step S3, the strategy for adjusting the water-cement ratio and the amount of early-strength agent in the explosive cartridge based on the tensile strength of the rock is as follows: when the tensile strength of the rock is ≤15MPa, the water-cement ratio is 0.1~0.3, and the amount of early-strength agent is 0.5% of the total mass of the explosive cartridge.

[0023] When 15MPa < rock tensile strength ≤ 25MPa, the water-cement ratio is 0.4~0.6, and the amount of early strength agent is 1.0% of the total mass of the cartridge.

[0024] When 25MPa < rock tensile strength < 35MPa, the water-cement ratio is 0.7~0.9, and the amount of early strength agent is 1.5% of the total mass of the cartridge.

[0025] When the tensile strength of the rock is ≥35MPa, the water-cement ratio is 1.0, and the amount of early strength agent is 2.0% of the total mass of the cartridge.

[0026] Preferably, in step S4, each segment of the drug cartridge is loaded into the borehole in order from the bottom to the top of the hole.

[0027] Preferably, in step S1, the drilled hole is a vertical hole with a diameter of 42~60mm.

[0028] Preferably, multiple rows of the boreholes are arranged along the direction of the tunnel, with a center-to-center spacing of 500-800 mm in the same row and a row spacing of 900-1200 mm.

[0029] The beneficial effects of this invention are as follows: First, the static fracturing method for roadway roof based on drilling detection provided by this invention integrates drilling parameters, P-wave velocity, and S-wave velocity, and uses a constructed rock uniaxial compressive strength inversion model and rock tensile strength analysis model for calculation. This provides a technical solution for obtaining rock tensile strength profiles in situ, rapidly, and continuously at the downhole construction site. This invention solves the technical problems of traditional methods that rely on core drilling and laboratory testing to obtain rock tensile strength, resulting in cumbersome procedures, long cycles, high costs, and susceptibility to rock sample disturbance. Furthermore, it shortens the acquisition cycle of key mechanical parameters from several days to be synchronized with drilling operations, thereby improving engineering efficiency and ensuring that the data accurately reflects the in-situ state of the rock strata.

[0030] Secondly, the static fracturing method for roadway roof based on drilling exploration provided by this invention divides the rock into lithological segments according to the tensile strength profile obtained by real-time inversion, and customizes static fracturing cartridges with different water-cement ratios and early strength agent dosages for each segment according to the pre-calibrated strength-ratio mapping relationship, thereby performing precise fracturing through "segment matching". In this way, this invention solves the technical problem that existing static fracturing processes rely on experience due to the lack of rock strength data, and the blind setting of parameters, which easily leads to fracturing failure or energy waste. Furthermore, it enables precise matching of fracturing energy with rock strength, promotes near-synchronous fracturing of different lithological segments and the formation of a through fracture network, thereby significantly improving the success rate of fracturing in one attempt and the overall weakening effect. Attached Figure Description

[0031] Figure 1 This is a schematic flowchart of a static fracturing method for roadway roof based on drilling detection, provided as an embodiment of the present invention.

[0032] Figure 2This is a schematic diagram showing the relationship between the uniaxial compressive strength of rock and the drilling depth in a static fracturing method for roadway roof based on drilling detection, provided as an embodiment of the present invention. Detailed Implementation

[0033] To make the above-mentioned objects, features and advantages of the present invention more apparent and understandable, the specific embodiments of the present invention will be described in detail below with reference to the accompanying drawings. Obviously, the described embodiments are only some embodiments of the present invention, and not all embodiments.

[0034] Example 1

[0035] This invention provides a static fracturing method for roadway roof based on drilling detection, the overall process of which can be found in [reference needed]. Figure 1 The following section will elaborate on the implementation methods, principles, logic, and beneficial effects of each part, following the step-by-step sequence.

[0036] S100, construction drilling and real-time acquisition of multi-source parameters.

[0037] As the data foundation for all subsequent analyses, the purpose of this step in the present invention is to simultaneously acquire multi-dimensional information flow reflecting the characteristics of the rock strata while forming a static fracturing physical channel (drill hole).

[0038] Specifically, conventional underground anchor bolt (cable) drilling rigs, such as pneumatic or hydraulic anchor bolt drilling rigs, are used to drill holes from the roadway floor or sidewalls into the roof strata. These holes are the fracturing holes for subsequent loading of static fracturing explosives. During the drilling process, two types of parameters need to be collected in real time and synchronously through a sensing system integrated on the drilling rig or drill rod.

[0039] Drilling mechanics parameters: These parameters directly reflect the mechanical processes of the interaction between the drill bit and the rock. They include: drilling rig propulsion force. The unit is kilonewton (kN), which refers to the force applied to the drill pipe to make it drill forward. It can be measured by a pressure sensor in the drilling rig's hydraulic system or a special pressure washer.

[0040] Drill bit speed The unit is revolutions per minute (r / min), which can be measured by an encoder to determine the rotational speed of the drive spindle.

[0041] Drilling torque The unit is Newton-meter (N·m), which refers to the torque required to rotate the drill bit against rock resistance. It can be measured by a torque sensor.

[0042] Drilling speed The unit is millimeters per minute (mm / min), which refers to the instantaneous advance rate of the drill bit in the drilling direction. It can be calculated by combining the displacement sensor with time.

[0043] Drilling depth The unit is meters (m), which refers to the current drilling depth from the borehole opening. It is usually obtained by integrating the drilling speed over time or by using a depth encoder.

[0044] Borehole acoustic parameters: These parameters reflect the elastic dynamics of the rock mass. During drilling, acoustic wave transmitters and receivers integrated near the drill bit or on the drill rod are used to excite and receive elastic waves in real time. By analyzing the received waveforms, the longitudinal wave velocity can be extracted. The unit is meters per second (m / s), which is the propagation speed of compression waves in rock mass. It is sensitive to the integrity and porosity of the rock mass.

[0045] transverse wave velocity The unit is meters per second (m / s), which is the propagation speed of shear waves in rock mass and is directly related to the shear modulus of the rock mass.

[0046] Therefore, this invention upgrades a typical drilling operation into a comprehensive in-situ testing project. By simultaneously acquiring mechanical and acoustic data, it provides dual information characterization of rock strata at the same depth point, including both "strength behavior" and "elastic response." This provides a more comprehensive and accurate description of the mechanical state of the rock strata compared to traditional single-type measurements, creating conditions for subsequent high-precision inversion. Furthermore, all the aforementioned data acquisition is completed simultaneously with the drilling operation, thereby maximizing work efficiency.

[0047] S200, Determination of rock tensile strength based on multi-source data.

[0048] This step transforms the raw, multi-dimensional, and noisy data collected by the S100 borehole into the direct basis for engineering decisions—the tensile strength of rock continuously distributed along the borehole depth. This step is further subdivided into four sub-steps: data processing, strength inversion, elastic parameter calculation, and comprehensive determination.

[0049] S210, Data Standardization and Dimensionality Reduction.

[0050] The raw collected parameters have different dimensions and orders of magnitude; therefore, standardization is required to eliminate the influence of dimensions. In this embodiment, the Z-score standardization method is used to process each type of parameter sequence separately, making its mean 0 and standard deviation 1.

[0051] Meanwhile, the standardized data still has a high dimensionality and may contain noise and redundancy. To extract core features and suppress noise, this invention employs the t-distributed random neighborhood embedding (t-SNE) algorithm for dimensionality reduction.

[0052] Specifically, t-SNE is an unsupervised machine learning algorithm for high-dimensional data visualization and feature extraction. Furthermore, this invention calculates the similarity (i.e., proximity probability) between data points in a high-dimensional space using a Gaussian distribution; then, in a low-dimensional space (such as two-dimensional or three-dimensional), it uses a t-distribution to represent this similarity, and iteratively optimizes it using gradient descent to minimize the KL divergence between the probability distributions of the high-dimensional and low-dimensional spaces. Ultimately, this ensures that similar points in the original high-dimensional space are closer to each other in the low-dimensional space, while dissimilar points are further apart. In this embodiment, the standardized multidimensional data is input into the t-SNE algorithm, and appropriate parameters such as perplexity are set. The final output is a dimensionality-reduced standard dataset that retains the main structure and features of the original data.

[0053] S220. Inversion of uniaxial compressive strength of rocks based on standard datasets.

[0054] Uniaxial compressive strength of rock is one of the most fundamental strength indicators of rock and is intrinsically related to drilling parameters. Based on drilling mechanics principles and extensive experimental data, this invention pre-constructs a dedicated inversion model for uniaxial compressive strength of rock. This model maps standardized and dimensionality-reduced feature data characterizing the current drilling state to the rock's UCS value.

[0055] In a preferred embodiment, the specific calculation formula for the inversion model is as follows: ;in, It represents the uniaxial compressive strength of the rock. For drilling rig propulsion, The linear attenuation coefficient of tensile stress is... The angle between the drill bit and the rock; The width of the drill bit's cutting edge. The rock cutting coefficient, The circumference of the drill bit. This is an empirical coefficient, usually taken as 0.1; For drilling speed, This represents the drill bit rotation speed.

[0056] Therefore, this invention can dynamically calculate the estimated value of the uniaxial compressive strength of the rock at the current drilling depth point based on real-time acquired drilling parameters, combined with known drill bit geometric parameters and empirical coefficients. As drilling progresses, a continuous curve showing the change of the uniaxial compressive strength of the rock with drilling depth can be obtained.

[0057] S230, Calculation of dynamic elastic parameters.

[0058] Specifically, this invention utilizes the longitudinal wave velocity (S100) collected by the device. ) and transverse wave velocity ( ), combined with rock density ( (This can be obtained through core sampling or assigned values ​​based on lithological experience), and two key dynamic elastic parameters are calculated: dynamic bulk modulus (…). The calculation formula is: Among them, the bulk modulus reflects the ability of a rock to resist volume changes under hydrostatic pressure.

[0059] Dynamic Poisson ratio ( The calculation formula is: Among them, the dynamic Poisson's ratio reflects the ratio of transverse strain to longitudinal strain of a rock under uniaxial stress.

[0060] S240, comprehensive determination of rock tensile strength.

[0061] Specifically, in this step, the present invention constructs a system that integrates the uniaxial compressive strength index (UCS) and the elastic index (dynamic bulk modulus) of rock. Dynamic Poisson ratio The rock tensile strength analysis model uses multiple regression fitting to achieve high-precision estimation of rock tensile strength.

[0062] In a preferred embodiment, the specific calculation formula of the analysis model is as follows: Among them, the fitting coefficients , and The values ​​are 0.01~0.02, 0.2~0.3, and 2~4, respectively; in this embodiment, the specific fitting coefficients are... , and The values ​​are 0.01, 0.2, and 2.4, respectively.

[0063] Therefore, by applying this model to the time series of rock uniaxial compressive strength, dynamic bulk modulus, and dynamic Poisson's ratio calculated point by point in S220 and S230, the present invention can finally output a high-precision curve of rock tensile strength as a function of drilling depth.

[0064] In summary, this invention, through a multi-step serial data processing and modeling process, "translates" the raw and mixed drilling and wave velocity signals into a key parameter—the rock tensile strength profile—that has direct guiding significance for static fracturing design, thereby improving the real-time performance and accuracy of the measurement.

[0065] S300, segmented custom static cracking explosive cartridge.

[0066] First, based on the rock tensile strength variation curve obtained from S240 with drilling depth, lithology segmentation is performed. The segmentation strategy is as follows: observe the rock tensile strength variation curve with drilling depth, and classify continuous segments on the curve where the fluctuation range of the rock tensile strength value is less than 3% into the same lithology segment. For example, from a depth of 1.0m to 2.5m, the rock tensile strength value fluctuates between 24.8MPa and 25.2MPa, with a maximum fluctuation range of (25.2-24.8) / 25.0≈1.6%<3%. Therefore, this 1.5m thick rock layer can be considered to have uniform tensile strength and belong to the same type of rock layer. This is then divided into one segment.

[0067] Record the segment number (a1, a2, ...), starting depth, segment length (l1, l2, ...), ending depth, and representative rock tensile strength of the segment (the average value of the segment can be taken).

[0068] Then, the customized production of the fracturing cartridges is carried out: Length customization: According to the segment lengths (l1, l2, ...) as defined above, the solid powdered static fracturing agent (mainly composed of CaO, etc.) is mixed with water in the corresponding proportion and stirred, then poured into a mold of a specific size, or directly packaged in a stretchable waterproof roll bag to produce solid static fracturing cartridges with lengths equal to the segment lengths (l1, l2, ...), and numbered accordingly as (b1, b2, ...).

[0069] In this embodiment, the formulation of the drug cartridge is adjusted according to the rock tensile strength of the rock layer corresponding to each segment of the drug cartridge and the optimal formulation strategy determined in advance through experiments: if the rock tensile strength of the segment is ≤15MPa, the water-cement ratio (water mass / dry powder mass) is controlled at 0.1 when preparing the drug cartridge, and an early strength agent (such as industrial calcium chloride) accounting for 0.5% of the total mass of dry powder is added to the dry powder.

[0070] If 15MPa < rock tensile strength ≤ 25MPa, then the water-cement ratio is 0.4 and the amount of early strength agent is 1.0%.

[0071] If 25MPa < rock tensile strength < 35MPa, then the water-cement ratio is 0.7 and the dosage of early strength agent is 1.5%.

[0072] If the tensile strength of the rock is ≥35MPa, the water-cement ratio is 1.0 and the dosage of early strength agent is 2.0%.

[0073] The water-cement ratio directly affects the hydration reaction rate and final peak expansion stress of the static fracturing agent. A lower water-cement ratio results in a thicker slurry, a more vigorous reaction, and greater expansion stress, but may shorten the reaction time. Early-strength agents can accelerate the hydration reaction process and shorten the time to reach peak stress. Therefore, this invention, based on experimental data, derives the above-mentioned correspondence, thus achieving a precise match between "high-energy, fast-reaction explosive cartridges for strong rocks and low-energy, slow-reaction explosive cartridges for weak rocks." This invention allows control over the expansion stress and hydration expansion time of each segment of the explosive cartridge, ensuring that all segments reach their peak expansion stress approximately synchronously within the borehole, with this peak slightly exceeding the tensile strength of the rock segment. Consequently, this invention ensures successful fracturing while avoiding energy waste and the problem of non-connected fractures caused by asynchronous fracturing.

[0074] S400, loading and sealing.

[0075] To facilitate operation and ensure accurate placement of the propellant cartridges, this invention employs an upward loading sequence from the bottom to the opening of the borehole. First, the propellant cartridge corresponding to the deepest segment is carefully pushed to the designed position at the bottom of the borehole using a delivery rod. Then, the propellant cartridges corresponding to the next higher rock layer are loaded sequentially until all segments fill their corresponding depth range. After loading, the borehole opening is immediately sealed with a specialized sealing material (such as fast-setting cement mortar or a specially designed sealing plug) to prevent leakage of propellant hydration products and ensure that expansion stress effectively acts on the borehole wall.

[0076] After the borehole is sealed, the customized explosive cartridges for each segment begin the hydration reaction. Due to the precise design of the formula, the expansion process of each cartridge is coordinated. As the reaction progresses fully, the expansion stress gradually increases and reaches its peak. When the circumferential tensile stress on the borehole wall rock exceeds its tensile strength, the rock cracks. Subsequently, different strength segments crack almost simultaneously, with cracks originating around each segment and extending outwards, eventually connecting to form a macroscopic fracture network penetrating multiple rock strata, thus effectively weakening the hard rock strata of the entire tunnel roof.

[0077] Because drilling needs to be planned for the entire fracturing zone, in this embodiment, the following drilling layout parameters are improved: each borehole is designed as a 90° vertical hole perpendicular to the roadway roof, and the hole diameter is selected as 45mm. This hole diameter can ensure sufficient charge and is also suitable for conventional anchor bolt drilling.

[0078] Multiple rows of fracture-causing holes are arranged along the direction of the tunnel (i.e., parallel to the tunnel axis) at a certain spacing (e.g., 1000mm).

[0079] Two fracture-inducing holes are arranged in the same row, with a spacing of 800 mm between them. This close-proximity arrangement of the two holes helps to create stress superposition between them, promoting fracture penetration.

[0080] Furthermore, during construction, following the procedures outlined in steps S100-S400, the drilling, exploration, analysis, explosive cartridge customization, filling, and sealing of each fracturing hole are completed sequentially. After all boreholes are drilled, the integrity and strength of the roof strata are significantly weakened by the synergistic fracturing effect of multiple rows and columns of boreholes, enabling it to collapse promptly and regularly under mine pressure, thereby effectively resolving safety hazards such as abnormal mine pressure and gas accumulation.

[0081]

Example 2

[0082] Unlike Example 1, the specific calculation formula for the inversion model provided in this example is as follows: In the formula: UCS represents the uniaxial compressive strength of rock; F is the drilling rig propulsion force, in kN; V is the drilling speed, in mm / min; N is the drill bit rotation speed, in r / min; k is the tensile stress linear attenuation coefficient, which is taken as 0.2 in this case; k1 is the rock cutting coefficient, which is taken as 0.25 here; β is the angle between the drill bit and the rock, which is taken as 90° here; b is the width of the drill bit cutting edge, which is taken as 14mm here; l is the circumference length of the drill bit, which is taken as 84mm here; μ is an empirical coefficient, which is taken as 0.1.

[0083] Collected longitudinal wave velocity ( ) and transverse wave velocity ( ), combined with rock density ( (This can be obtained through core sampling or assigned values ​​based on lithological experience), and two key dynamic elastic parameters are calculated: dynamic bulk modulus (…). The calculation formula is: Among them, the bulk modulus reflects the ability of a rock to resist volume changes under hydrostatic pressure.

[0084] Dynamic Poisson ratio ( The calculation formula is: Among them, the dynamic Poisson's ratio reflects the ratio of transverse strain to longitudinal strain of a rock under uniaxial stress.

[0085] The specific calculation formula for this analytical model is as follows: .

[0086] In this embodiment, the fitting coefficient , and The values ​​are 0.015, 0.245, and 3.7. The calculation results based on the above equations are shown in the appendix. Figure 2 Taking the example shown, the rock tensile strength corresponding to different depth ranges of statically induced fracture holes can be obtained. The total length of the statically induced fracture holes is 10m. Through inversion of drilling data, it can be seen that within the range of 0.00~1.25m (… The tensile strength of the limestone in this segment is approximately 22.35 MPa, and the length of this segment is... It is 1.25m.

[0087] Within the range of 1.25~3.52m ( The tensile strength of the siltstone in this segment is approximately 17.58 MPa, and the length of this segment is... It is 2.27m.

[0088] Within the range of 3.52~7.19m ( The tensile strength of the granite segment is approximately 25.03 MPa, and the length of this segment is... It is 3.57m.

[0089] Within the range of 7.19~10.00m ( The tensile strength of the sandstone in this segment is approximately 19.82 MPa, and the length of this segment is... It is 2.91m.

[0090] Based on the above segmentation results, the solid static fracturing explosive cartridge filled in the static fracturing hole is divided into 4 segments, numbered b1, b2, b3, and b4, with lengths of 1.25m, 2.27m, 3.57m, and 2.91m, respectively.

[0091] Adjust the water-cement ratio and early strength agent dosage of the static cracking agent cartridge according to the uniaxial tensile strength of the rock, and control the expansion stress and hydration expansion time of each segment of the cartridge. The selection principle of the parameters of each segment of the static cracking agent cartridge is as follows: ① When the tensile strength of the rock is ≤15MPa, the water-cement ratio is 0.3 and the dosage of early strength agent is 0.5% of the total mass of the cartridge.

[0092] ② When 15MPa < rock tensile strength ≤ 25MPa, the water-cement ratio is 0.5, and the amount of early strength agent is 1.0% of the total mass of the cartridge.

[0093] ③ When 25MPa < rock tensile strength < 35MPa, the water-cement ratio is 0.8, and the amount of early strength agent is 1.5% of the total mass of the cartridge.

[0094] ④ When the tensile strength of the rock is ≥35MPa, the water-cement ratio is 1.0, and the amount of early strength agent is 2.0% of the total mass of the cartridge.

[0095] Based on the above principles, the technical parameters of solid static cracking explosive cartridges b1, b2, b3, and b4 can be determined and are summarized in Table 1 below.

[0096] Table 1: Technical parameters of static crack-inducing solid explosive cartridges for each segment

[0097]

[0098] The static fracturing hole constructed here has a diameter of 50mm and is drilled vertically at an azimuth angle of 90°.

[0099] The prepared static fracturing solid explosive cartridges are loaded into the static fracturing holes in the order of b4, b3, b2, b1, and then sealed. As the hydration reaction of each segment of the static fracturing explosive cartridge gradually becomes sufficient and the expansion stress reaches its peak, the rock around the hole cracks synchronously and forms a through fracture, ultimately achieving segmented and multiple static fracturing of the hard overlying rock of the tunnel roof.

[0100] Two static fracturing holes are arranged in a single row along the tunnel direction, with a hole spacing of 600mm.

[0101] Multiple rows of static fracturing holes are arranged sequentially along the roadway, with a hole spacing of 900mm, and static fracturing is carried out in sequence.

[0102] Those skilled in the art will understand that embodiments of the present invention can be provided as methods, systems, or computer program products. Therefore, the present invention can take the form of a completely hardware embodiment, a completely software embodiment, or an embodiment combining software and hardware aspects. Furthermore, the present invention can take the form of a computer program product implemented on one or more computer-usable storage media containing computer-usable program code. The storage medium can be implemented by any type of volatile or non-volatile storage device or a combination thereof, such as Static Random Access Memory (SRAM), Electrically Erasable Programmable Read-Only Memory (EEPROM), Erasable Programmable Read-Only Memory (EPROM), Programmable Read-Only Memory (PROM), Read-Only Memory (ROM), magnetic storage, flash memory, magnetic disk, or optical disk. These computer program instructions may also be stored in a computer-readable storage medium that can direct a computer or other programmable data processing device to function in a particular manner, such that the instructions stored in the computer-readable storage medium produce an article of manufacture including instruction means, which are implemented in a process Figure 1 One or more processes and / or boxes Figure 1 The function specified in one or more boxes.

[0103] It should be noted that the above embodiments are only used to illustrate the technical solutions of the present invention and are not intended to limit it. Although the present invention has been described in detail with reference to preferred embodiments, those skilled in the art should understand that modifications or equivalent substitutions can be made to the technical solutions of the present invention without departing from the spirit and scope of the technical solutions of the present invention, and all such modifications or substitutions should be covered within the scope of the claims of the present invention.

Claims

1. A method for static fracturing of roadway roof based on drilling detection, characterized in that, Includes the following steps: S1. Drilling into the roof of the tunnel and collecting drilling parameters, as well as longitudinal and transverse wave velocities in the borehole, in real time during the drilling process. The drilling parameters include drilling rig thrust, drill bit speed, drilling torque, drilling speed and drilling depth. S2. Based on the drilling parameters, the longitudinal wave velocity, and the transverse wave velocity, determine the rock tensile strength corresponding to different depth ranges of the borehole; S3. Based on the different depth ranges and their corresponding rock tensile strengths, prepare solid static fracturing explosive cartridges that match the length of each depth range, and adjust the water-cement ratio and early strength agent dosage of the corresponding explosive cartridges according to the rock tensile strength of each depth range to control the expansion stress and hydration expansion time of each segment of the explosive cartridge. S4. Load the adjusted segmented explosive cartridges into the corresponding depth range of the drill hole and seal the hole; Step S2 includes: S21. The collected drilling parameters, P-wave velocity, and S-wave velocity are standardized and dimension-reduced to obtain a standard dataset. S22. Based on the standard dataset and the pre-built rock uniaxial compressive strength inversion model, determine the rock uniaxial compressive strength along the depth direction of the borehole; S23. Based on the longitudinal wave velocity and the transverse wave velocity, calculate the dynamic bulk modulus and dynamic Poisson's ratio of the rock. S24. Based on the uniaxial compressive strength of the rock, the dynamic bulk modulus, and the dynamic Poisson's ratio, determine the tensile strength of the rock using a pre-constructed rock tensile strength analysis model; In step S24, the dynamic bulk modulus The calculation formula is as follows: ； in, The density of the rock can be obtained by core sampling or assigned a value based on lithological experience. For longitudinal wave velocity, The transverse wave velocity; The dynamic Poisson ratio The calculation formula is as follows: ； The calculation formula for the rock tensile strength analysis model is as follows: ； in, , and The fitting coefficients were obtained through fitting analysis of a large amount of indoor experimental data; UCS represents the uniaxial compressive strength of the rock. In step S2, the strategy for dividing the different depth ranges is as follows: Based on the curve of the rock tensile strength as a function of drilling depth, continuous segments with curve amplitude fluctuations of less than 3% are divided into the same depth range; In step S3, the strategy for adjusting the water-cement ratio and the amount of early-strength agent in the injection cartridge based on the tensile strength of the rock is as follows: When the tensile strength of the rock is ≤15 MPa, the water-cement ratio is 0.1~0.3, and the dosage of the early strength agent is 0.5% of the total mass of the cartridge. When 15 MPa < rock tensile strength ≤ 25 MPa, the water-cement ratio is 0.4~0.6, and the dosage of early strength agent is 1.0% of the total mass of the cartridge; When 25 MPa < rock tensile strength < 35 MPa, the water-cement ratio is 0.7~0.9, and the dosage of early strength agent is 1.5% of the total mass of the cartridge; When the tensile strength of the rock is ≥35 MPa, the water-cement ratio is 1.0, and the amount of early strength agent is 2.0% of the total mass of the cartridge.

2. The method for static fracturing of roadway roof based on drilling detection according to claim 1, characterized in that, In step S21, the t-SNE algorithm is used to reduce the dimensionality of the data.

3. The method for static fracturing of roadway roof based on drilling detection according to claim 1, characterized in that, In step S4, the segmented cartridges are loaded into the borehole in order from the bottom to the top.

4. The method for static fracturing of roadway roof based on drilling detection according to claim 1, characterized in that, In step S1, the drilled hole is a vertical hole with a diameter of 42~60mm.

5. The method for static fracturing of roadway roof based on drilling detection according to claim 1, characterized in that, Multiple rows of the aforementioned boreholes are arranged along the direction of the tunnel, with a center-to-center spacing of 500-800mm within the same row and a row spacing of 900-1200mm.

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