A stability control method for replacing primary pillars with artificial pillars in a phosphate ore goaf

Abstract

The application discloses a kind of stability control method of phosphate ore goaf artificial pillar replacement primary pillar, belong to the technical field of goaf management in mine.The method is first by field sampling, indoor rock mechanics test and drilling detection, obtain the physical and mechanical parameters of surrounding rock and ore body and rock mass quality index, adopt three-level grading system to complete rock mass quality refinement grading, form double-track parameter;Again, calculate the safety factor of primary pillar based on the above parameters, determine the initial stability state of goaf, and combine contribution area method and double-threshold safety system, quantitatively design the cross-sectional size and material parameters of artificial pillar;Then construct a three-dimensional stope numerical model, complete model mechanical parameter assignment by multi-criteria quantitative mapping, simulate the whole replacement process and verify the safety of the scheme.The application realizes the quantitative stability control of the whole replacement process, effectively guarantees the construction and operation safety, improves the reliability and efficiency of goaf management engineering.

Description

Technical Field

[0001] This invention belongs to the field of underground mining and goaf management technology, specifically relating to a stability control method for replacing primary pillars with artificial pillars in phosphate mine goaf areas. Background Technology

[0002] Phosphate rock is a vital natural resource, characterized by its non-renewable nature. Underground phosphate mining is constrained by various factors, including production processes, environment, and technical conditions, leading to widespread phenomena such as "mining the rich and abandoning the poor" and "mining above ground and leaving the rest." This disorderly mining not only results in a significant waste of mineral resources but also leaves behind large areas of mined-out areas, posing substantial safety hazards to mine production. Therefore, safe and efficient phosphate mining has become a critical problem that needs to be solved.

[0003] Full-scale mining, as one of the main methods of underground phosphate mining, presents significant risks to mine safety, particularly in the goaf. Pillars, as the sole supporting structure within the goaf in full-scale mining, are the location where stress transfer and elastic potential energy converge. If a pillar becomes unstable, the overlying load transfers to adjacent pillars, causing them to become overloaded and ultimately unstable. This phenomenon, known as "domino" pillar instability, can easily lead to overlying strata collapse, surface subsidence, and severe mine pressure.

[0004] The use of artificial concrete pillars combined with full-scale mining can effectively maintain the stability of the stope structure during mining, and has the advantages of high recovery rate and low dilution rate. However, this technology faces a core challenge in its implementation: the stability control of the surrounding rock in the goaf during the replacement process. The removal of the original pillars disrupts the original stress balance. If the newly constructed artificial pillars are insufficient in strength, have unreasonable dimensions, or are constructed at an inappropriate time, it can easily lead to roof instability, chain reactions of pillar failure, or even surface collapse. Existing technologies rely heavily on experience-based design, and each stage (rock mass evaluation, pillar design, numerical simulation, and field monitoring) is independent. There is a lack of a systematic and coordinated control method that directly uses the results of refined rock mass classification to calibrate numerical model parameters and uses monitoring data to close the loop and correct design parameters, making it difficult to ensure precise control and long-term safety of the replacement process. Summary of the Invention

[0005] The purpose of this invention is to provide a stability control method for replacing primary pillars with artificial pillars in phosphate mine goaf areas, addressing the problems of isolated technical links, distorted rock mass parameter transmission, and disconnect between design and monitoring in existing technologies. The core innovation of this invention lies in: establishing for the first time a closed-loop parameter transmission chain from RMR, IRMR, MRMR, rock mass parameter mapping, numerical model, and monitoring back-calibration; converting rock mass quality classification results using the Hoek-Brown criterion and Mohr-Coulomb, quantifying and mapping them into the strength parameters required by the numerical model; and using field monitoring data to verify the design safety factor in a closed loop, thus solving the problems of parameter fragmentation and large deviations in empirical values ​​in traditional methods.

[0006] To achieve the above objectives, the present invention adopts the following technical solution:

[0007] A method for stability control in replacing primary pillars with artificial pillars in phosphate mine goaf areas, the method comprising the following steps:

[0008] Step S1: On-site sampling and indoor rock mechanics tests are conducted on the phosphate mine to be replaced to obtain the physical and mechanical parameters of the surrounding rock and ore body; rock mass quality indicators are obtained through borehole exploration; a three-level grading system of rock mass geomechanical grading - improved rock mass geomechanical grading - corrected rock mass geomechanical grading is adopted to complete the fine grading of rock mass quality, forming a dual-track parameter of corrected rock mass geomechanical grading deterioration score and original rock mass geomechanical grading score;

[0009] Step S2: Using the physical and mechanical parameters as the basis for calculating the stability of the primary pillar and designing the parameters of the artificial pillar, calculate the safety factor of the primary pillar and determine the initial stability state of the goaf; based on the contribution area method and the dual threshold safety system, quantitatively design the cross-sectional dimensions and material parameters of the artificial pillar.

[0010] Step S3: Based on the dual-track parameters, the cross-sectional dimensions and material parameters of the quantitatively designed artificial pillar, a three-dimensional stope numerical model is constructed. The mechanical parameters of the three-dimensional stope numerical model are assigned by quantitative mapping through rock mass geomechanical classification, geological strength index, Hawke-Brown strength criterion, and Mohr-Coulomb strength criterion. The entire replacement process is simulated and the safety of the scheme is verified by quantitative criteria. If the conditions are not met, the process returns to step S2 for iterative optimization.

[0011] Preferably, the method further includes: step S4, implementing artificial pillar casting and primary pillar replacement operations according to the scheme verified in step S3, deploying monitoring equipment to collect full-cycle stress and displacement data, and performing closed-loop verification of the monitoring data with the design parameters of step S2 and the simulation results of step S3 to complete the stability control of the entire replacement process.

[0012] Preferably, in step S1, the abnormal mechanical parameters obtained from the indoor rock mechanics test are reduced, and parameter range sensitivity analysis is carried out to avoid numerical model distortion caused by parameter anomalies.

[0013] Preferably, in step S1, the three-level classification system of rock mass geomechanical classification - improved rock mass geomechanical classification - corrected rock mass geomechanical classification adopts a dual-track division of labor mechanism:

[0014] First, the preliminary score of the rock mass geomechanical classification is calculated. Then, the improved rock mass geomechanical classification score is obtained after size effect correction. Finally, the improved rock mass geomechanical classification deterioration score is obtained by multiplying the five correction coefficients together.

[0015] The original score of the rock mass geomechanical classification is used for subsequent parameter mapping, while the modified rock mass geomechanical classification deterioration score is only used for early warning of surrounding rock stability.

[0016] Preferably, in step S2, the safety factor of the primary pillar is calculated using the Binyawski pillar strength formula, and the contribution area of ​​the pillar roof is divided using the area apportionment method. When the safety factors of all primary pillars are ≥1.5, the goaf is determined to meet the initial stability conditions for replacement.

[0017] Preferably, in step S2, when quantitatively designing the cross-sectional dimensions of the artificial pillar, the minimum cross-sectional area of ​​the artificial pillar is quantitatively calculated using the following formula:

[0018] ;

[0019] in, This represents the minimum cross-sectional area of ​​an artificial pillar. A safety factor is designed for short-term use; This is a dynamic load correction factor; This refers to the vertical stress of the overlying rock strata. Contributes area to the top of the pillar; The effective strength of the artificial pillar in the field.

[0020] Preferably, in step S2, the dual-threshold safety system is as follows: for the stage of artificial pillar replacement construction and mining impact, a safety factor of 1.5 is used for bearing stability verification; for the stage of stable operation of the goaf, after introducing a creep reduction coefficient to reduce the effective strength of the artificial pillar, a safety factor of 1.3 is used for bearing stability verification.

[0021] Preferably, in step S3, the quantitative mapping process is as follows: the original score of the rock mass geomechanical classification is converted into a geological strength index, the rock mass strength parameters are calculated by combining the Hawke-Brown strength criterion constant and the disturbance factor, and then converted into the Mohr-Coulomb strength criterion parameters by the tangent method. The artificial pillar is assigned values ​​using independent concrete mechanical parameters.

[0022] Preferably, in step S4, when carrying out the artificial pillar casting and primary pillar replacement operations, an interval arrangement and skip mining replacement construction strategy is adopted: the pillars are grouped into odd and even groups, the odd-numbered artificial pillars are constructed first and the corresponding primary pillars are removed after curing to the required standard, and then the even-numbered artificial pillars are constructed and the corresponding primary pillars are removed after curing to the required standard. Adjacent primary pillars are not removed at the same time.

[0023] Preferably, in step S4, when setting up monitoring equipment, collecting data, and performing closed-loop verification with design parameters and simulation results, the following closed-loop monitoring operations are performed: stress gauges are installed inside the artificial pillar to collect stress data every 10 days, and stress gauges and delamination instruments are installed on the roof of the goaf to collect displacement data every 7 days; the safety factor and simulation accuracy are verified with measured data.

[0024] The present invention provides a stability control method for replacing primary pillars with artificial pillars in phosphate mine goaf areas, which achieves several technical advantages:

[0025] (1) A closed-loop collaborative control system of classification-design-simulation-monitoring was established: This invention establishes for the first time a full-chain quantitative method from RMR / IRMR / MRMR classification, GSI conversion, Hoek-Brown / Mohr-Coulomb parameter mapping, numerical model, design parameters, and monitoring back-verification, which solves the problem of parameter fragmentation in traditional methods where rock mass classification information is only used for qualitative description and cannot be quantitatively transferred to subsequent design stages. This closed-loop link makes the output of each technical stage become the quantitative input of the next stage. Combined with the closed-loop verification of monitoring data, the dynamic reliability verification of design parameters during construction and service is realized.

[0026] (2) Full-process quantitative control has been achieved: This invention transforms the stability criteria from subjective qualitative descriptions such as no obvious delamination and allowable displacement into measurable and verifiable quantitative standards such as roof displacement ≤ 1 / 500 of the mining area span, plastic zone penetration rate ≤ 30%, and stress safety factor ≥ 1.5, so that the safety evaluation has authority and reproducibility.

[0027] (3) Improved design economy and safety: This invention, through precise quantitative design, can optimize the size of artificial pillars while ensuring safety, maximizing resource recovery rate. The pillars designed using C25 concrete with an effective strength of 9.5 MPa have a measured maximum stress of only 5.76 MPa, a short-term stress safety factor of approximately 1.65, and a long-term check safety factor of approximately 1.40, both meeting their respective threshold requirements. The approximately 18% deviation between the numerical simulation maximum principal compressive stress of 4.87 MPa and the measured 5.76 MPa falls within the reasonable error range between model simplification and the actual rock mass mechanical response. The design uses measured data as the final control basis, further ensuring the reliability of the safety assessment. Through meticulous construction and full-process monitoring, the risks of the replacement project are controlled within the preset acceptable range.

[0028] (4) The construction process is operable and safe: This invention discloses key procedures such as the on-site layered casting construction method for artificial pillars, the curing period, the construction sequence strategy of interval arrangement and skip mining replacement, and the timing and method of removing primary pillars. The cross-sectional design of the artificial pillar is based on... =9.5MPa uniform calculation. Due to the large area required, KZ-3 and KZ-4 adopt a segmented pillar scheme, replacing a single continuous pillar with two segmented pillars to solve the space adaptation problem and ensure the passage and ventilation needs of loaders in the mine. The control principle of not removing adjacent original pillars at the same time ensures the continuity of roof stress and provides a clear operational guide for engineering practice. Attached Figure Description

[0029] Figure 1 The flowchart shows steps S1-S3 of a stability control method for replacing primary pillars with artificial pillars in a phosphate mine goaf according to the present invention.

[0030] Figure 2 The results of drilling and inspection of the left side of the horizontal tunnel for segmented transportation.

[0031] Figure 3 The results of drilling into the roof of the horizontal tunnel for segmented transportation.

[0032] Figure 4 The results of drilling and inspection of the right side of the horizontal tunnel for segmented transportation.

[0033] Figure 5 This is a stress cloud diagram after the mining operation is completed.

[0034] Figure 6 The flowchart shows steps S1-S4 of a stability control method for replacing primary pillars with artificial pillars in a phosphate mine goaf according to the present invention.

[0035] Figure 7 This is the stress-time curve of an artificial pillar. Detailed Implementation

[0036] The following detailed implementation of a method for stability control of artificial pillars replacing primary pillars in phosphate mine goaf areas, based on specific embodiments, is provided below. These embodiments are for illustrative purposes only and are not intended to limit the scope of protection of this invention.

[0037] Example 1: Implementation of a stability control method for replacing primary pillars with artificial pillars in phosphate mine goaf areas.

[0038] Combined with appendix Figures 1-5 As shown, the present invention provides a stability control method for replacing primary pillars with artificial pillars in phosphate mine goaf areas.

[0039] like Figure 1 As shown, Figure 1 This is a flowchart of steps S1-S3 of a stability control method for replacing primary pillars with artificial pillars in a phosphate mine goaf according to Embodiment 1 of the present invention.

[0040] Step S1: Comprehensive investigation and evaluation of the engineering geology and rock mechanics properties of the goaf.

[0041] In the replacement mining area, the system acquires the physical and mechanical parameters of the surrounding rock and ore body, and performs refined rock mass quality classification, directly providing quantitative input parameters for subsequent numerical models and designs.

[0042] Step S1 involves on-site sampling and indoor rock mechanics tests at the phosphate mine to be replaced, obtaining the physical and mechanical parameters of the surrounding rock and ore body; rock mass quality indicators are obtained through borehole exploration, and a three-level grading system of rock mass geomechanical grading, improved rock mass geomechanical grading, and corrected rock mass geomechanical grading is adopted to complete the refined grading of rock mass quality, forming a dual-track parameter of the corrected rock mass geomechanical grading deterioration score and the original rock mass geomechanical grading score.

[0043] In step S1, the abnormal mechanical parameters obtained from the indoor rock mechanics test are reduced, and parameter range sensitivity analysis is carried out to avoid numerical model distortion caused by parameter anomalies.

[0044] In step S1, the three-tiered rock mass geomechanical classification system—improved rock mass geomechanical classification, and revised rock mass geomechanical classification—adopts a dual-track division of labor mechanism: first, a preliminary rock mass geomechanical classification score is calculated; then, after size effect correction, an improved rock mass geomechanical classification score is obtained; finally, a revised rock mass geomechanical classification deterioration score is obtained by multiplying five correction coefficients together. The original rock mass geomechanical classification score is used for subsequent parameter mapping, while the revised rock mass geomechanical classification deterioration score is only used for early warning of surrounding rock stability. The specific process is as follows:

[0045] Step S1.1, Determination of rock mass mechanical parameters:

[0046] By taking samples on site and conducting indoor rock mechanics tests, the uniaxial compressive strength, tensile strength, cohesion, internal friction angle, elastic modulus, and Poisson's ratio of the rocks in the ore body, roof, and floor were obtained.

[0047] The rock sample collection and preparation process is as follows: To ensure the accuracy and reliability of rock mechanics testing, for the No. 2 phosphate orebody and its roof (dolomite) and floor (green shale) of the North 2#778 mining area of ​​Ping'an Phosphate Mine No. 2, the rock samples collected on-site need to be transported to the laboratory and processed into standard specimens before the mechanical parameter testing. Specifically, this includes preparing 50mm×100mm uniaxial compressive strength standard specimens, 50mm×50mm shear test standard specimens, and 50mm×25mm tensile test (Brazilian splitting method) standard specimens, with 4 specimens for each type of lithology, for a total of 36 standard specimens.

[0048] To ensure the accuracy and reliability of the test data, the geometric accuracy of the specimen processing must be strictly controlled: both end faces must be ground flat to ensure that their parallelism meets the requirements—no more than 0.02 mm for uniaxial compressive strength test and no more than 0.25 mm for Brazilian splitting test; the perpendicularity of the end face to the specimen axis must also be strictly controlled, no more than 0.01 rad for uniaxial test (or no more than 0.05 mm deviation within a length of 50 mm), and no more than 0.25° for Brazilian splitting test; in addition, the unevenness of the specimen side surface should not exceed 0.3 mm for uniaxial test and no more than 0.025 mm for unevenness in the thickness direction for Brazilian splitting test.

[0049] The uniaxial compressive strength test was conducted using a static universal testing machine and a matching strain data acquisition instrument. The specimen was loaded at a rate of 0.5–1 MPa / s until failure, and the failure load and phenomena observed during loading were recorded. The uniaxial compressive strength was calculated using the following formula:

[0050] ;

[0051] In the formula: The uniaxial compressive strength (MPa) of the rock. The maximum failure load of the rock specimen is kN; The area of ​​the specimen subjected to pressure. .

[0052] elastic modulus Defined as the ratio of stress to corresponding strain in a rock mass under uniaxial stress: In the formula The stress borne by the rock. ; This represents the linear strain produced in a rock under stress.

[0053] Poisson's ratio Defined as the ratio of transverse strain to axial strain in a rock mass under uniaxial tension or compression: In the formula For lateral strain, For axial strain.

[0054] The shear strength test employed the variable-angle shear test method, performing shearing at four different angles: 45°, 50°, 55°, and 60°. The normal stress and shear stress on the specimen were calculated using the following formula:

[0055] ;

[0056] In the formula: The peak shear load is (kN). The effective shear area (m²) Let be the normal stress on the shear surface, in MPa; The shear stress on the shear surface is expressed in MPa. Let be the tilt angle of the fixture, respectively. , The coefficient of friction between the clamping balls and the pressure plate is taken as 0.

[0057] Using normal stress and shear stress data from four angles, in The fitted curve in the rectangular coordinate system has its ordinate intercept corresponding to the cohesion of the rock mass. The slope corresponds to the friction angle within the rock mass. The tangent value is used to obtain the cohesion of the rock. and internal friction angle .

[0058] The tensile strength test was conducted using the Brazilian splitting method, with a loading rate of 0.5-1 MPa / s until the specimen failed. The tensile strength was calculated using the following formula:

[0059] ;

[0060] In the formula: The tensile strength of the rock mass is given in MPa. The peak load at which the specimen fails. ; For the diameter of the specimen ; The height of the specimen is denoted as mm; the average value of the four sets of uniaxial tensile strengths is taken as the final uniaxial tensile strength of the rock mass.

[0061] The unit weight of rock is determined by the volumetric method and calculated using the following formula:

[0062] ;

[0063] In the formula: The natural density of the rock. ; The volume of the rock sample being tested. ; The mass of the rock sample being tested is expressed in kg. Let be the acceleration due to gravity, taken as 9.8.

[0064] The following rock mechanical parameters were obtained through experiments. Table 1 shows the test results of the uniaxial compressive strength of the rock mass.

[0065] Table 1

[0066] Top plate dolomite DY-1 49.29 100.14 79.60 72.30 22.42 0.29 Top plate dolomite DY-2 49.47 100.08 67.41 Top plate dolomite DY-3 49.32 100.30 75.83 Top plate dolomite DY-4 49.50 99.96 66.37 Phosphate rock KY-1 49.76 100.43 49.70 47.72 13.58 0.33 Phosphate rock KY-2 49.81 100.18 51.27 Phosphate rock KY-3 49.79 100.19 48.76 Phosphate rock KY-4 49.76 100.27 41.15 Green shale base dY-1 49.24 99.72 55.28 66.38 11.71 0.22 Green shale base dY-2 49.32 100.08 66.63 Green shale base dY-3 49.21 100.11 79.86 Green shale base dY-4 49.28 99.94 63.75

[0067] Table 2 shows the test results of cohesion and internal friction angle of the rock specimens.

[0068] Table 2

[0069] Dolomite 9.29 40.16 Phosphate rock 7.97 40.64 green shale 17.65 24.14

[0070] Explanation of Data Reasonableness: The cohesion of the bottom green shale is as high as 17.65 MPa, significantly higher than that of general shale, which is usually 1-8 MPa, and also higher than that of the top dolomite in this mining area, at 9.29 MPa. After verifying the original test data and conducting on-site geological surveys, the geological origin of this abnormally high value is as follows: The bottom green shale in this mining area belongs to the Doushantuo Formation of the Lower Sinian System, with a lithology of grayish-green fine to medium-grained quartz sandstone. The lower part contains an increased amount of argillaceous material and generally contains scattered euhedral pyrite grains, indicating strong diagenetic compaction and siliceous cementation. However, considering that this high value deviates from the distribution pattern of conventional rock mechanics parameters, and that the combination of uniaxial tensile strength (9.19 MPa) and low friction angle (φ=24.14°) is atypical within the Mohr-Coulomb framework, the following measures are taken in the engineering design to avoid the risk of model distortion caused by single-point anomalies:

[0071] (1) Design value reduction: The design value of cohesion Cd of green shale is reduced to 0.7 times the test value, i.e., Cd=17.65×0.7≈12.4MPa, in order to cover the adverse effects of test dispersion and rock mass heterogeneity.

[0072] (2) Interval sensitivity analysis: In the numerical simulation, the sensitivity of the cohesion parameter of the bottom green shale was verified in the range of 10-18 MPa. The analysis results show that the distribution and displacement of the plastic zone of the surrounding rock in the goaf are not sensitive to the change of C value within this range, and the roof stability evaluation conclusion remains consistent, further verifying the reliability of the design scheme.

[0073] With the above conservative treatment, the bearing capacity of the surrounding rock will not be underestimated or its stability will not be overestimated in subsequent numerical simulations and engineering designs.

[0074] Table 3 shows the test results of the tensile strength of the rock samples.

[0075] Table 3

[0076] Table 4 shows the results of the rock bulk density test.

[0077] Table 4

[0078] In summary, the physical and mechanical parameters of the ore and rock in the North 2#778 stope of Ping'an Phosphate Mine No. 2 are summarized as follows (the cohesion and friction angle listed in the table are values ​​determined by indoor direct shear tests. The surrounding rock parameters actually used in the numerical model are the equivalent rock mass parameters transformed by the Hoek-Brown criterion, and the artificial pillars are independently parameterized using C25 concrete):

[0079] Table 5 shows the physical and mechanical parameters of the ore and rock strata (values ​​determined by indoor tests).

[0080] Table 5

[0081] Dolomite 2770 17.79 8.69 0.29 9.29 40.16 10.67 Ore body No. II ore layer 2700 13.31 5.10 0.33 7.97 40.64 4.49 Green shale 2740 6.97 4.80 0.22 12.4* 24.14 9.19 sandstone 2520 18.24 10.24 0.23 20.24 45 3.52 filling body 1890 0.13 0.10 0.20 0.20 35.20 0.5

[0082] The cohesion of the green shale is the design value after a reduction of 0.7 times; the original experimental value is 17.65 MPa.

[0083] Explanation of sandstone parameter sources: The physical and mechanical parameters of the sandstone layer in Table 5, located beneath the bottom green shale, were determined by referring to the measured data of the lower quartz sandstone of the Doushantuo Formation of the same Sinian system provided in the "Detailed Exploration Report of the Expansion of Ping'an Phosphate Mine No. 2 in Jinzhong Town, Kaiyang County, Guizhou Province," and by combining the sampling test results of sandstone of the same stratum in adjacent mining areas. This sandstone layer is located at the bottom of the model, with a vertical distance of more than 100m from the goaf of the phosphate mine. Numerical simulation sensitivity analysis shows that within a reasonable fluctuation range of sandstone parameters, such as an elastic modulus of 15–25 GPa and a cohesion of 15–25 MPa, the variation range of the displacement and plastic zone of the surrounding rock in the goaf does not exceed 3%, and its influence can be ignored.

[0084] Step S1.2, rock mass quality classification and integrity evaluation.

[0085] A borehole inspection instrument was used to detect the development of joints and fractures in the roof and sidewalls of the stope, and to obtain rock mass quality indicators (RQD). For example... Figure 2-4 As shown, Figure 2 The results of the drilling inspection of the left side of the horizontal tunnel in the segmented transportation are shown in Figures (a)-(f), which show the inspection results of the drilling at 0.5m, 2m, 3m, 5m, 6m and 7.5m respectively. Figure 3 The results of drilling into the roof of the segmented transport tunnel are shown in Figures (a)-(f), which show the observations of the roof drilling at 0.5m, 1.5m, 3m, 4.5m, 6m, and 8m, respectively. Figure 4The results of the drilling on the right side of the segmented transport tunnel are shown in Figures (a)-(f), which show the drilling conditions at 0.5m, 2m, 3m, 4m, 6m, and 7.8m respectively.

[0086] Drilling inspection tests were conducted in the sectional transport roadway of the North 2#778 mining area. The vertical boreholes had a diameter of 32mm, a depth of 8m, and a spacing of 10m. The inclined boreholes were located 1.5m below the roof, drilled at a 45° angle, with a diameter of 32mm and a depth of 8m.

[0087] Drilling revealed the following characteristics: On the left side of the segmented haulage tunnel, at a depth of 0.5m, the rock mass was smooth, intact, and dry. At a depth of 2m, a disturbance zone with increased roughness appeared. In the depth range of 3m to 8m, the borehole wall returned to a smooth state, with no joints or fissures, and remained dry, indicating good rock mass integrity and structural stability in this section. In the roof of the segmented haulage tunnel, at a depth of 0.5m, a locally fractured area extending approximately 10cm was observed. At a depth of 1.5m, the borehole wall exhibited a rough morphology. In the depth range of 2m to 8m, the borehole wall was relatively smooth, with no obvious joints or fissures, indicating that the tunnel roof at depths deeper than 2m possessed good rock integrity and structural stability. On the right side of the segmented transport tunnel, the integrity of the rock mass structure in the 0.5m to 3m section is further enhanced; when the borehole reaches a depth of 4m, the borehole wall is wet due to local groundwater seepage; in the 6m to 8m depth section, the borehole wall returns to a dry state, and the surface is smooth, with an intact rock mass structure.

[0088] Based on the statistical results of intact core segments with a length ≥10cm, the RQD values ​​of each lithology were calculated as follows: Table 6 shows the RQD values ​​and rock quality grades of the North 2#778 mining area.

[0089] Table 6

[0090] Dolomite 81.2 II good rock mass Relatively complete Phosphite 76.9 II good rock mass Relatively complete green shale 72.1 III Medium-sized rock mass Medium complete

[0091] A comprehensive evaluation of rock mass quality was conducted using RMR, IRMR, and MRMR grading methods.

[0092] First, preliminary classification is performed using the RMR method. Rock Geomechanical Classification (RMR) is a quantitative evaluation system based on the physical and mechanical properties and structural characteristics of rock masses, proposed by Z.T. Bieniawski in 1976. The RMR method calculates the RMR value of rock masses using six indicators: rock block strength, RQD value, joint spacing, joint state, groundwater conditions, and joint direction and dip angle, through a cumulative scoring method, thus achieving a quantitative classification of rock mass quality. The RMR indicator value standards are as follows: Table 7 shows the RMR indicator value standards.

[0093] Table 7

[0094] Intact rock strength / MPa ≥250 (15 minutes) / 100-250 (12 minutes) / 50-100 (7 minutes) / 25-50 (4 minutes) / 5-25 (2 minutes) / <5 (0 minutes) Rock quality index RQD / % 90-100 (20 minutes) / 75-90 (17 minutes) / 50-75 (13 minutes) / 25-50 (8 minutes) / <25 (3 minutes) Joint spacing / cm ≥200 (20 minutes) / 60-200 (15 minutes) / 20-60 (10 minutes) / 6-20 (8 minutes) / <6 (5 minutes) Joint state Very rough, discontinuous, closed and unweathered (30 points) / Rough, opening <1mm, slightly weathered hard rock wall (25 points) / Slightly rough, opening <1mm, slightly weathered soft rock wall (20 points) / Mirror-like or mud inclusion thickness <5mm, or opening 1-5mm, continuous joints (10 points) / Mud inclusion thickness >5mm or opening >5mm, continuous joints (0 points) groundwater status Completely dry (15 points) / Damp (10 points) / Wet (7 points) / Can form water droplets (4 points) / Can form water streams (0 points)

[0095] The unit for joint spacing is cm, consistent with the original RMR grading standard of Bieniawski (1976). In the joint condition score, the key difference between 25 and 20 points is that 25 points corresponds to hard rock face, with high joint surface roughness and discontinuous joints; 20 points corresponds to soft rock or medium-strength rock face, with slightly rougher joint surfaces and higher continuity.

[0096] The RMR correction scoring table based on joint occurrence is as follows: Table 8 is the RMR correction scoring table based on joint occurrence.

[0097] Table 8

[0098] score 0 -2 -5 -10 -12

[0099] The RMR calculation results are as follows: Table 9 shows the RMR calculation results and rock mass classification.

[0100] Table 9

[0101] roof 72.30 (7 points) 81.2 (17 points) 32.7 (10 points) Rough, opening <1mm, hard rock wall (25 points) Wet (7 points) Average (-5 points) 61 (Level II) Ore body 47.72 (4 points) 76.9 (17 points) 22.1 (10 points) Slightly rough, opening <1mm, medium-hard rock face, high continuity (20 points) Damp (10 points) Favorable (-2 points) 59 (Level III) base plate 66.38 (7 points) 72.1 (13 points) 29.5 (10 points) Slightly rough, opening <1mm, soft rock wall (20 points) Damp (10 points) Average (-5 points) 55 (Level III)

[0102] Then, the IRMR method is introduced. The IRMR (Improved Geomechanical Classification of Rock Mass) classification system was established by DH Laubscher et al. Based on the RMR method, the core evaluation indicators are rock strength (RBS), joint spacing (JS), and joint state (JC). However, it does not fully consider the key factor of joint attitude, and it needs to be corrected according to the degree of influence of joint attitude after obtaining the preliminary RMR value.

[0103] The IRMR score is calculated using the following formula: IRMR = RBS + JS + JC.

[0104] In the formula: RBS is the rock block strength score, JS is the joint spacing score, and JC is the joint state evaluation value.

[0105] The above three scoring items all directly adopt the corresponding scoring criteria in the RMR system, as detailed in Table 7. The specific correspondence is as follows:

[0106] (1) RBS (Rock Block Strength Score): Corresponds to the "Intact Rock Strength / MPa" row in Table 7. The score is based on the measured uniaxial compressive strength value in S1.1, according to the standards in Table 7. For example, the compressive strength of the dolomite roof is 72.30 MPa, falling within the 50–100 MPa range, corresponding to a score of 7. In the IRMR system, the RBS score needs further consideration of size effect correction (multiplied by a correction factor of 0.8), resulting in RBS = 7 × 0.8 = 5.6 points.

[0107] (2) JS (Joint Spacing Score): Corresponds to the "Joint Spacing / cm" row in Table 7. The joint spacing data obtained from borehole inspection and on-site measurement are scored according to the standards in Table 7. For example, if the measured joint spacing of the top plate is 32.7cm, which falls within the range of 20-60cm, the corresponding score is 10 points.

[0108] (3) JC (Joint State Score): Corresponds to the "Joint State" row in Table 7. Based on the joint surface roughness, opening, filling condition and weathering degree observed by borehole inspection, the score is calculated according to the standards in Table 7. For example, a rough joint surface with an opening of less than 1 mm, slight weathering, and hard rock wall corresponds to a score of 25 points.

[0109] Finally, the MRMR method is used to further correct for factors such as weathering degree, blasting damage, and stress changes caused by mining, and the rock mass quality is finally corrected in combination with the actual mining conditions. The MRMR (Modified Geomechanical Classification of Rock Mass) system was developed by Laubscher based on IRMR. It not only considers the influence of initial stress and disturbance stress, but also takes into account the weakening of rock mass properties caused by blasting and weathering.

[0110] The MRMR rock mass evaluation method multiplies the IRMR score by a series of correction factors, and the calculation formula is as follows: .

[0111] The signs, meanings, and values ​​of each correction coefficient in the formula are explained below:

[0112] In Laubscher's MRMR system, the size effect correction is already reflected in the RBS score of IRMR; the cemented structure surface correction is already reflected in the JC score of IRMR. Therefore, in the comprehensive correction stage of MRMR, only five factors need to be corrected: weathering, occurrence, mining activity, blasting, and groundwater, i.e., five coefficients from A3 to A7 are involved in the multiplication.

[0113] The specific values ​​of each correction factor are as follows: Table 10 shows the rock mass correction factors.

[0114] Table 10

[0115] For a conservative assessment in engineering, the mining stress correction factor A5 is set to 1.00, which means that the local strength enhancement effect that may be caused by the redistribution of mining stress is not considered, and the value is taken on the safe side.

[0116] Taking the roof as an example, the complete MRMR calculation process is as follows:

[0117] Step 1: Calculate IRMR;

[0118] RBS = 7 (original strength score) × 0.8 (size correction) = 5.6;

[0119] JS=10 (Joint spacing score, directly adopting the standard in Table 7);

[0120] JC=25 (Joint status score, directly adopting the standard in Table 7);

[0121] IRMR=RBS+JS+JC=5.6+10+25=40.6;

[0122] Step 2: Calculate the MRMR correction factor multiplication;

[0123]

[0124] Step 3: Calculate MRMR;

[0125] MRMR=IRMR×0.8722=40.6×0.8722=35.4;

[0126] The product of the above correction coefficients is 0.8722 < 1, which is consistent with the essential characteristics of the MRMR system as a deterioration correction system. The corrected rock mass score is lower than the initial score, reflecting the comprehensive weakening effect of factors such as weathering, blasting, joint occurrence and groundwater on the quality of the rock mass.

[0127] The final rock mass rating results are as follows: Table 11 shows the calculation results and rock mass rating.

[0128] Table 11

[0129] roof 5.6 10 25 40.6 35.4 Grade IV, Poor Phosphate deposits 3.2 10 20 33.2 30.9 Grade IV, Poor base plate 5.6 10 20 35.6 27.9 Grade IV, Poor

[0130] The rock mass classification standards are as follows: Table 12 shows the rock mass classification standards.

[0131] Table 12

[0132] The above results indicate that the roof, phosphate ore body, and floor are all rated as Grade IV, indicating poor rock mass. This reflects that mining disturbances and other engineering activities significantly reduced the stability of the surrounding rock. Therefore, to ensure the safety of subsequent mining operations, appropriate reinforcement and control measures must be taken for the stope roof and pillar system.

[0133] In this invention, the MRMR method and the subsequent GSI (geological strength index) conversion method serve different purposes and have their own roles in the parameter transfer chain. The MRMR method comprehensively considers engineering disturbance factors such as weathering, blasting, mining stress, and groundwater, and is used to assess the actual stability level of the surrounding rock in the mining area under the current mining conditions to determine whether reinforcement and control measures are needed. On the other hand, the geological strength index GSI in the Hoek-Brown criterion (i.e., the Hoek-Brown strength criterion) characterizes the inherent geological properties of the rock mass itself, such as the degree of development of structural planes and rock block strength. In the conversion, the original RMR score value without engineering disturbance correction should be used, because disturbance factors such as mining and blasting have been uniformly considered in the disturbance factor D of the subsequent numerical model and should not be repeatedly reduced in the parameter conversion. This dual-track parallel and complementary design is one of the core methodologies of the closed-loop parameter transfer chain of this invention. MRMR provides stability early warning and reinforcement necessity criteria, while RMR, GSI, Hoek-Brown, and Mohr-Coulomb (i.e., Mohr-Coulomb strength criteria) provide the geological basis parameters of the numerical model. Together, they serve the safety and reliability of artificial pillar design.

[0134] Step S2: Stability analysis of primary pillars and quantitative design of artificial pillar parameters.

[0135] Under the premise of confirming that the current state of the goaf is stable, the precise design of artificial pillars is carried out.

[0136] Step S2 uses the physical and mechanical parameters as the basis for calculating the stability of the primary pillar and designing the parameters of the artificial pillar. It calculates the safety factor of the primary pillar and determines the initial stability state of the goaf. Based on the contribution area method and the dual threshold safety system, it quantitatively designs the cross-sectional dimensions and material parameters of the artificial pillar.

[0137] In step S2, the safety factor of the primary pillar is calculated using the Binyawski pillar strength formula. The contribution area of ​​the pillar roof is divided using the area apportionment method. When the safety factors of all primary pillars are ≥1.5, the goaf is determined to meet the initial stability conditions for replacement.

[0138] The dual-threshold safety system described in step S2 is as follows: For the stage involving artificial pillar replacement construction and mining impact, a safety factor of 1.5 is used for bearing stability verification; for the stage of stable operation of the goaf, a creep reduction factor is introduced to reduce the effective strength of the artificial pillar, and then a safety factor of 1.3 is used for bearing stability verification. The specific process is as follows:

[0139] Step S2.1, Quantitative analysis of the stability of the primary ore pillar:

[0140] The safety factor for all primary pillars was calculated using the Bieniawski pillar strength formula.

[0141] Pillar strength formula:

[0142] ;

[0143] Safety factor formula:

[0144] ;

[0145] in: The strength of the pillar is expressed in MPa. The uniaxial compressive strength of the pillar rock, in MPa, was obtained through indoor rock mechanics tests in step S1.1. For the northern... Phosphate ore from the mine, ; The width of the pillar is in meters (m). The height of the pillar is in meters (m). This is an empirical index related to the aspect ratio of the pillar and rock properties. Based on the original suggestions of Bieniawski (1975) and the systematic discussion by Bieniawski ZT in *Engineering Rock Mass Classifications* (John Wiley & Sons, 1989), The rules for determining the value are as follows: (1) When the width-to-height ratio of the pillar is... hour, Take 1.0, which is suitable for short and thick pillars; (2) When the width-to-height ratio of the pillar is 1.0, it is suitable for short and thick pillars; hour, A value of 0.5 is suitable for slender pillars. North The aspect ratio of each primary pillar in the stope is within Within the range (see Table 13), all values ​​are less than 1, therefore, the calculation is in progress. Let's all take 0.5. conditions, The range of values ​​for the term is approximately This reflects the strength reduction effect of slender pillars relative to standard cubic specimens; The safety factor for the pillar; The unit weight of the overlying rock strata ( ), ; Thickness of the overlying rock layer , ; The area of ​​the roof supported by a single pillar. ; Let be the cross-sectional area of ​​the pillar. .

[0146] The area supported by the pillars is divided according to the method of area apportionment. That is, for a group of pillars, the area supported by the pillars is divided by the midpoint connecting the adjacent pillars. If the pillar is adjacent to the boundary of the goaf, the roof area to be apportioned is determined by the line connecting the midpoints of the two pillars.

[0147] The pillars in the North 2#778 stope are numbered sequentially from the innermost side outwards. The measurement results and safety factor calculations are as follows: Table 13 shows the pillar safety factor calculation based on Bieniawski.

[0148] Table 13

[0149] KZ-1 13.2 5.2 6.5 0.80 250.50 68.64 4.04 stable KZ-2 13.6 5.2 6.5 0.80 366.00 70.72 2.85 stable KZ-3 13.1 6.1 6.5 0.94 403.50 79.91 3.08 stable KZ-4 8.9 5.3 6.5 0.82 372.75 47.17 1.88 stable KZ-5 12 5.4 6.5 0.83 363.75 64.80 2.66 stable KZ-6 11.8 5.4 6.5 0.83 356.25 63.72 2.67 stable

[0150] To ensure long-term stability of the goaf, the safety factor of the pillar must reach or exceed 1.5. According to the data in the table above, the minimum safety factor for the goaf pillar is 1.88, and all are greater than 1.5, indicating that the current goaf is in a stable state. Here, the safety factor of 1.5 refers to the long-term stability factor for the primary pillar.

[0151] Step S2.2, Minimum size design of artificial pillars.

[0152] The artificial pillars numbered kz-1 to kz-6 below correspond to the primary pillars KZ-1 to KZ-6 in Table 13, respectively, indicating that the primary pillars were replaced by artificial pillars at the same location.

[0153] The design is based on the contribution area method and the safety factor method.

[0154] First, the material and effective strength of the artificial pillars must be determined. In full-scale mining operations, artificial pillars are crucial for maintaining stope stability and ensuring mining safety. Artificial pillars are supporting structures built in the goaf, and commonly used materials include concrete and waste rock binders. Compared to natural pillars, the size, shape, and specific location of artificial pillars can be flexibly adjusted and set according to the actual needs of the mining site.

[0155] The artificial mine pillar material is C25 concrete. The standard value of the 28-day cubic compressive strength of C25 concrete is... Its axial compressive strength design value is based on the "Code for Design of Concrete Structures" (GB 50010). Considering the construction quality, geometric defects, long-term effects of a humid environment, and the influence of the pillar's height-to-width ratio, a comprehensive reduction factor of 0.8 is used to characterize the effective strength in the field, yielding the effective strength. The reduction factor of 0.8 is composed as follows: a long-term strength reduction factor for underground humid environments of approximately 0.85; and a reduction factor for pillar size effect and construction deviation of approximately 0.94. This factor is based on the stability coefficient table for axially compressed members in Appendix D of GB 50010, corresponding to the ratio of the calculated length of the pillar to the radius of gyration (slenderness ratio). The stability coefficient of ). The product of the two. .

[0156] The design adopts a dual-threshold safety system: the short-term load-bearing safety factor is 1.5 (construction period and early mining stage), and the long-term service safety factor is 1.3 (operation period, referring to the provisions on long-term load reduction of permanent support structures in relevant specifications such as the "Technical Specification for Rock Anchors and Shotcrete Support Engineering" GB50086).

[0157] The vertical self-weight stress at the burial depth is calculated using the following formula:

[0158] ;

[0159] Pick , ,but .

[0160] This value is a theoretical calculation based on the self-weight stress of the overlying strata. In the subsequent numerical model in section S3.1, considering the arching and unloading effect of the goaf roof, the influence of the lateral pressure coefficient, and the difference between the free boundary at the top of the model and the actual infinite half-space, the applied equivalent vertical load is reduced. The actual applied value is 16.42 MPa, and the reduction factor is approximately 0.92.

[0161] Let the contribution area of ​​a single pillar be Considering the effects of arching, filling sharing, and stage replacement, a load correction factor is introduced. Then the axial force on the pillar is:

[0162] ;

[0163] Design stability coefficient This is a short-term design safety factor specifically for the bearing capacity of artificial mine pillars, covering material strength variability and construction deviations. For the long-term service phase (operational period), a creep reduction factor is introduced. It is then verified in conjunction with a long-term safety factor threshold of 1.3. The following conditions must be met:

[0164] ;

[0165] The minimum cross-sectional area of ​​the artificial pillar is obtained by sorting out:

[0166] ;

[0167] in: This is the minimum cross-sectional area required for an artificial pillar. ; For short-term design safety factor, take ; This is the dynamic load correction factor during the replacement period, with a value range of [value missing]. . The determination is based on the additional dynamic loads that may occur during the replacement construction: when the original pillar is removed by static cutting or mechanical means. Take 1.0; when blasting operations are involved, Take 1.2 ; The vertical stress of the overlying strata at the location of the pillar, in MPa. ; The area contributing to the ore pillar. ; The effective strength of the artificial pillar material is taken as 9.5 MPa.

[0168] If represented by an equivalent square prism, then the equivalent side length The equivalent side length is assumed only to be the minimum side length when using a square cross-section, serving as a geometric feature reference. Actual designs use rectangular cross-sections, and the verification is based on the design area. Required area, not shorter side length Equivalent side length.

[0169] Contribution area of ​​pillars in North 2#778 mining area This refers to the supporting area of ​​each primary pillar in Table 13. Taking the KZ-1 to KZ-6 pillars of the North 2#778 mining area as an example, during the replacement phase... Take 1.0 and verify with kz-1: Similarly, the calculation results for the minimum cross-sectional requirements of artificial pillars are as follows: Table 14 shows the minimum cross-sectional requirements of artificial pillars under C25 conditions.

[0170] Table 14

[0171] The equivalent side length is the minimum side length assumed to be a square cross-section and is only used for geometric reference. The actual design uses a rectangular cross-section, with the length direction consistent with the original ore pillar (along the ore body strike), and the width calculated in reverse based on the required area. The verification basis for the cross-section design is that the design area ≥ the required area, not that the shorter side length ≥ the equivalent side length. In the calculation, the design value of the axial compressive strength of C25 concrete is taken as 11.9 MPa (GB 50010), and the effective strength is taken as 9.5 MPa (with a comprehensive reduction factor of 0.8).

[0172] Due to the large area required for KZ-3 and KZ-4, and the fact that the width of a single pillar would exceed the passage space limitations of the minehouse, a segmented pillar scheme was adopted in actual construction: two segmented pillars were used instead of a single continuous pillar. KZ-3: Single pillar cross-section 13.1×4.6m, area 60.3m², two pillars total 120.6m², meeting the required area of ​​113.84m², with a passage gap of at least 2m between pillars, and approximately 3.7m of clearance between each pillar and the minehouse boundary on both sides. KZ-4: Single pillar cross-section 8.9×6.0m, area 53.4m², two pillars total 106.8m², meeting the required area of ​​105.18m², with a passage gap of at least 2m between pillars, and approximately 3m of clearance between each pillar and the minehouse boundary on both sides. Both schemes meet the passage and ventilation requirements of the loader. Based on this, the final design dimensions and layout of the artificial pillars were determined.

[0173] Step S3: Numerical simulation analysis and effect verification of the entire replacement process.

[0174] Before formal construction, a detailed three-dimensional numerical model is established to simulate the replacement process.

[0175] Step S3, based on the dual-track parameters, the cross-sectional dimensions and material parameters of the quantitatively designed artificial pillar, constructs a three-dimensional stope numerical model. The mechanical parameters of the three-dimensional stope numerical model are assigned through the quantitative mapping of rock mass geomechanical classification, geological strength index, Hawke-Brown strength criterion, and Mohr-Coulomb strength criterion. The entire replacement process is simulated and the safety of the scheme is verified by quantitative criteria. If the conditions are not met, the process returns to step S2 for iterative optimization.

[0176] The quantitative mapping process in step S3 is as follows: the original rock mass geomechanical classification score is converted into a geological strength index; the rock mass strength parameters are calculated by combining the Hawke-Brown strength criterion constant and the disturbance factor; then, the parameters are equivalently converted to Mohr-Coulomb strength criterion parameters using the tangent method; and the artificial pillars are assigned values ​​using independent concrete mechanical parameters. The specific process is as follows:

[0177] Step S3.1, Model Establishment.

[0178] Based on the actual geological conditions and mining layout, a three-dimensional finite difference numerical model was established, which includes the roof, ore body, floor, primary pillars, and proposed artificial pillars.

[0179] The three-dimensional finite difference calculation software FLAC3D, developed by Itasca Consulting Group Inc. in the United States, was used for three-dimensional mechanical calculation and analysis of the ore body mining process. The FLAC3D program has a variety of constitutive models, including isotropic elastic material models, Mohr-Coulomb elastoplastic material models, strain softening / hardening plastic material models, empty element models, etc., which can be used to simulate the excavation of underground chambers and the mining of ore bodies.

[0180] Taking the North 2#778 stope as an example, the model dimensions are length × width × height = 289m × 133m × 150m, divided into 745,136 units. The model's strata, from top to bottom, consist of dolomite on top, a phosphate rock layer, and green shale and sandstone on the bottom. The X-direction of the model represents the dip of the ore layer, and the Y-direction represents its strike. The average thickness of the phosphate rock layer is 3.5m, with an average dip angle of 25°. Two segmented stopes are arranged within the simulation area: North 2#778 and North 2#770. The lower part of North 2#770 is a backfill, and the goaf is supported by concrete pillars. The North 2#778 goaf retains its original pillars for pre-replacement stability assessment, or replaces them with artificial pillars for post-replacement stability assessment.

[0181] The model adopts the Mohr-Coulomb (i.e., Mohr-Coulomb strength criterion) elastoplastic constitutive model, which is one of the most commonly used constitutive models in geotechnical engineering and can well describe the shear failure characteristics of rock materials under compressive stress.

[0182] The model's X and Y direction boundaries are subject to normal displacement constraints, meaning the normal displacement is zero. The bottom boundary is fixed, with zero displacement in all three directions. The top boundary is a free surface with an equivalent overlying rock load of 16.42 MPa. This 16.42 MPa is the engineering application value of the theoretical self-weight stress of 17.88 MPa after reduction by the arching effect and lateral pressure coefficient, and is consistent with the design load value for the artificial pillar in Section S2.2. The mesh is generated using hexahedral elements, with local refinement in the pillar and roof regions. The element size in the refined areas ranges from 0.5 to 2.0 m.

[0183] Mapping method from rock mass classification results to numerical model parameters: The surrounding rock mechanical parameters input to the numerical model are quantitatively converted from the RMR rock mass score obtained in step S1.2 through the following steps:

[0184] (1) RMR to GSI conversion: For rock masses with RMR>23, the geological strength index GSI is calculated using the following formula: GSI=RMR-5. The original RMR score is used here instead of the corrected MRMR value because GSI represents the inherent geological properties of the rock mass itself rather than the state of engineering disturbance. Accordingly, the RMR=61 to GSI=56 for the top dolomite; the RMR=59 to GSI=54 for the phosphate ore body; and the RMR=55 to GSI=50 for the bottom green shale.

[0185] (2) Hoek-Brown Criterion Parameter Calculation: Based on the GSI value, the generalized Hoek-Brown strength criterion was used to determine the rock mass strength parameters. The disturbance factor D was set to 0.5 to uniformly consider the influence of blasting damage and mining-induced stress relaxation on the rock mass strength. These two factors were evaluated in the MRMR correction stage, and this is reflected in the Hoek-Brown parameters through the D factor to avoid repeated reduction. Hoek-Brown Parameters Calculate using the following formula:

[0186] ;

[0187] In the formula The Hoek-Brown constant represents the intact rock. (Various lithologies) The values ​​are as follows: 9 for dolomite, which conforms to the value table for dolomite in Hoek & Brown (1997). Reference value; for phosphate rock, take 12, referring to limestone in sedimentary rocks. The range is set to a conservative, higher value to reflect the high strength characteristics of the dense phosphorite in this mining area; the green shale, actually sandy shale / sandy mudstone, is set to 8. The reason for choosing 8 is: according to the geological description of this rock layer in step S1.1: the lithology is grayish-green fine to medium-grained quartz sandstone with increased argillaceous content. This rock layer belongs to a transitional type between shale and sandstone in the Hoek-Brown classification system. Typical shale... The reference value is 4-6, while that of typical sandstone is... The reference value is 17-19, with sandy shale serving as an intermediate transitional type between the two. exist A value between these ranges is reasonable. In this design, [the value is taken as...]. This value is consistent with the value suggested by Marinos & Hoek (2001) for sandy shale. The range is 8-12, with the lower limit consistent, and it is more conservative than typical shale, avoiding the influence of... The value was too low, which led to a serious underestimation of the rock mass strength.

[0188] (2) Equivalent Mohr-Coulomb parameter transformation: Based on the Hoek-Brown parameters, the tangent method is used to equivalently transform the cohesion required for the Mohr-Coulomb model. and friction angle This is used for assigning values ​​to FLAC3D models. During tangent method conversion, the confining pressure range is taken. ,Right now At a pressure of 18.1 MPa, the best-fit tangent to the Hoek-Brown envelope is obtained within this range, yielding the equivalent cohesion and friction angle. The conversion results are shown in the table below.

[0189]

[0190] Green shale C test The values ​​are the original values ​​from the indoor tests, and 12.4 MPa is a 0.7 times reduction of the design value; the numerical model finally adopts the equivalent rock mass parameters after Hoek-Brown transformation.

[0191] It should be noted that the cohesion and friction angle listed in Table 5 are parameters of intact rock determined by indoor direct shear tests. The numerical model actually uses the equivalent rock mass scale parameters transformed by the Hoek-Brown criterion, i.e., the "Values ​​Used in Numerical Model" column in the table above. The difference between the two sets of parameters reflects the size effect and joint reduction from intact rock to engineering rock mass. The remaining parameters in Table 5, such as density, elastic modulus, and Poisson's ratio, continue to use their original values ​​in the numerical model.

[0192] Independent parameter assignment for artificial mine pillars (C25 concrete) in the FLAC3D model: The artificial mine pillar element in the model is assigned independent elastoplastic parameters to C25 concrete, which are not confused with the filling parameters in Table 5. The values ​​assigned to the concrete in the FLAC3D model are: density 2400 kg / m³, elastic modulus E = 28 GPa, referring to the elastic modulus of C25 concrete in GB 50010, bulk modulus K = E / [3(1-2ν)] = 15.56 GPa, shear modulus G = E / [2(1+ν)] = 11.67 GPa, Poisson's ratio ν = 0.2, cohesion c = 5.0 MPa, concrete cohesion is converted from axial compressive strength of 11.9 MPa according to the Mohr-Coulomb criterion, friction angle φ = 35°, and tensile strength 1.78 MPa. The above parameters are based on the standard values ​​for C25 concrete in GB 50010, and are determined in conjunction with the reduction factor for the humid environment in underground mines. In step S3.2, the safety factor is calculated using the effective concrete strength of 9.5 MPa, and the stress safety factor is calculated by comparing it with the maximum principal compressive stress output by the model.

[0193] Feasibility study of construction space: According to the mine design data, the cross-section of the sectional horizontal roadway in the North 2#778 mining area is an irregularly shaped roadway of 4.2 m × 3.5 m. The stope is divided into blocks along the strike at intervals of 30-50 m, and the net width of the stope is approximately 12-15 m. The design cross-section of the artificial pillars in Table 14 is based on... The spatial feasibility analysis for each pillar is as follows:

[0194] (1) kz-1, kz-2, kz-5, kz-6: The design widths are 5.6m, 7.8m, 8.8m and 8.8m respectively. In a 12m net width minehouse, the remaining widths on both sides of the mine pillar are approximately 6.4m, 4.2m, 3.2m and 3.2m respectively. After deducting a safety passage distance of not less than 2m, there is still a margin of more than 1.2m to meet the passage requirements of the loader.

[0195] (2) KZ-3: The required area is 113.84 m². If a single column with a cross-section of 13.1 × 9.0 m and a width of 9.0 m is used, only 1.5 m of clearance remains on both sides within the 12 m net width of the minehouse, which does not meet the passage requirements. In actual construction, a segmented pillar scheme is adopted: two segmented pillars are used instead, with a single column cross-section of 13.1 × 4.6 m and an area of ​​60.3 m². The two pillars together cover 120.6 m², with a passage gap of no less than 2 m between the pillars. Approximately 3.7 m is reserved on both sides of the pillars and the boundary of the minehouse, which meets the passage requirements.

[0196] (3) KZ-4: The required area is 105.18 m², and the original pillar length is only 8.9 m. If a single pillar with a cross-section of 8.9 × 12.0 m and a width of 12.0 m is adopted, it is close to the net width of the mine. In actual construction, a segmented pillar scheme is adopted: two segmented pillars are used instead, with a single pillar cross-section of 8.9 × 6.0 m and an area of ​​53.4 m². The two pillars together cover 106.8 m². A passage gap of no less than 2 m is left between the pillars, and about 3 m is reserved on both sides of the pillars and the boundary of the mine to meet the passage requirements.

[0197] Step S3.2, simulation analysis.

[0198] The entire process of artificial pillars replacing primary pillars was simulated, and the stress field, displacement field, and plastic zone distribution characteristics of the surrounding rock in the goaf were analyzed. For example... Figure 5 As shown, Figure 5 Figure 1 shows the stress cloud diagram after the mining is completed. Figure 2 shows the maximum principal stress of the roof, Figure 3 shows the minimum principal stress of the roof, Figure 4 shows the maximum principal stress of the pillar, and Figure 5 shows the minimum principal stress of the pillar.

[0199] There are three main criteria for judging whether a goaf is stable: first, whether the stress on the artificial pillars and their roof exceeds the rock strength; second, whether the artificial pillars and roof exceed their limit displacement; and third, whether the artificial pillars and roof show signs of plastic zone failure.

[0200] Stability is determined based on the following quantitative criteria:

[0201] (1) The maximum displacement of the roof does not exceed 1 / 500 of the short span of the mining area, which is about 0.3m in this embodiment. The short span of the mining area is 15m.

[0202] (2) The stress safety factor of the artificial pillar, that is, the ratio of the effective concrete strength of 9.5 MPa to the maximum principal compressive stress output by the model is not less than 1.5.

[0203] (3) The plastic zone penetration rate, that is, the percentage of the plastic zone area to the cross-sectional area of ​​the pillar does not exceed 30%, and the roof does not form a penetrating plastic zone.

[0204] (4) Verify whether the artificial pillar is under the main stress.

[0205] Simulation results of goaf stability before replacement under the original pillar condition:

[0206] The roof stress cloud diagram shows that the maximum principal stress ranges from 4.22 MPa to -5.19 MPa, exhibiting characteristics of both tensile and compressive stress. Locally occurring tensile stress zones are mostly distributed in the middle of the goaf or in the roof bending and delamination areas; widely distributed compressive stress zones are concentrated at the edges of the goaf, above the pillar supports, and at the junction of the roof and surrounding rock. The minimum principal stress ranges from 0.012 MPa to -43.33 MPa, indicating an overall strong triaxial compression state.

[0207] The stress state of the pillar exhibits a more pronounced pressure concentration characteristic. Its maximum principal stress ranges from 4.22 MPa to -5.70 MPa; the minimum principal stresses are all compressive, ranging from -0.0058 MPa to -27.95 MPa, with significant high-pressure stress cores forming particularly in the middle of the pillar and in the contact area with the roof and floor. Comparing this maximum compressive stress with the ultimate compressive strength of the pillar rock (47.72 MPa), the maximum compressive stress does not exceed its ultimate compressive strength.

[0208] The roof displacement cloud map shows significant spatial heterogeneity: the maximum positive displacement is 1.32 cm, mainly distributed in the middle of the goaf; the maximum negative displacement reaches -1.74 cm, appearing at the edge of the goaf or in the area affected by the pillar support. The pillar displacement cloud map shows an overall negative displacement, with values ​​ranging from 0 to -8.59 mm.

[0209] The cloud map of the plastic zone of the roof shows that tensile failure plastic zone is mainly developed in the central part of the goaf and the decompression zone; while shear failure and combined shear failure are the main characteristics at the edge of the goaf, above the pillar support, and at the junction of the roof and the two sides. The pillar as a whole is dominated by shear failure and combined shear failure, but the degree of damage in the entire goaf is relatively small.

[0210] The above results indicate that the goaf is in a stable state under the original pillar condition. The pillars are able to withstand the overburden pressure of the roof and maintain the stability of the roof and pillars by virtue of their own strength.

[0211] Simulation results of goaf stability after artificial pillar replacement:

[0212] The maximum principal stress cloud map of the replaced roof section shows values ​​ranging from approximately 5.16 MPa to -5.16 MPa, exhibiting a significant alternating distribution of positive and negative values. The minimum principal stress cloud map of the roof section ranges from approximately +0.324 MPa to -28.02 MPa. The maximum principal stress cloud map of the artificial pillar section shows values ​​ranging from approximately +4.87 MPa to -3.85 MPa, also exhibiting a coexistence of tensile and compressive stresses. The minimum principal stresses of this section are all compressive stresses, ranging from -0.59 MPa to -20.91 MPa. The fact that all stresses are compressive stresses is consistent with the stress characteristics of an artificial pillar as a core supporting component, indicating that it is under three-dimensional compression, which is conducive to utilizing the compressive strength of the material.

[0213] The displacement contour plot of the top slab after replacement shows that the displacement ranges from 2.3120 × 10⁻⁶. -3 m to -2.9042×10 -3 The maximum uplift displacement is approximately 2.31 mm, and the maximum settlement displacement is approximately -2.90 mm. The displacement contour map of the artificial pillar profile shows a displacement range between 2.3120 × 10⁻⁶ m. -3 m to -4.3922×10 -4 The displacement ranges from m. The maximum negative displacement is approximately -0.44 mm, indicating that the main body of the pillar undergoes primarily compressive settlement deformation, consistent with the mechanical behavior of the pillar under the vertical load of the overlying strata. The small positive displacement, approximately 2.31 mm, is only distributed in the contact area between the pillar bottom and the floor slab, mainly due to the rebound effect caused by unloading during surrounding rock excavation and boundary effects in numerical calculations, and not an overall uplift of the pillar. Overall, the artificial pillar is in a stable compressive state, playing a significant vertical supporting role and effectively constraining the downward deformation of the roof slab.

[0214] The plastic zones in the replaced roof profile were relatively localized, with no large-scale continuous failure observed. The localized tensile plastic zones suggest a risk of tensile damage in the central or bending areas of the roof, while the scattered shear plastic zones were mostly located at the contact edges between the pillar and the roof. The artificial pillar profile showed a clear concentration of shear plastic zones, indicating that the pillar, as the core load-bearing structure, had entered a shear yielding state, and its interior and the contact area with the roof and floor were subjected to high compressive and shear stresses.

[0215] Quantitative simulation results show that the maximum displacement of the roof after replacement is approximately 2.9 mm, far less than 1 / 500 of the short-span of 15 m, i.e., 30 mm. The maximum principal compressive stress of the artificial pillar in the numerical model is approximately 4.87 MPa, corresponding to a stress safety factor of 9.5 / 4.87≈1.95, which is greater than the short-term threshold of 1.5. No continuous plastic zone was formed in the roof; the plastic zone exhibits a locally discontinuous distribution. All quantitative indicators meet safety standards. The replacement project achieved spatial transfer of the risk of plastic failure: roof failure was effectively controlled, while the artificial pillar became the main bearer of plastic deformation, absorbing energy and regulating stress through its own yielding, thus keeping the goaf in a stable and controllable state.

[0216] Step S3.3, the plan is determined.

[0217] If the simulation results do not meet any of the above quantitative indicators, then return to S2.2, adjust the size and strength of the artificial mine pillar, or modify the replacement sequence, and recalculate and simulate until all safety requirements are met. In this embodiment, the C25 concrete artificial mine pillar is... =9.5MPa. The required area was calculated and the cross-section was designed. The simulation results met all safety standards, and the scheme was determined.

[0218] Example 2, as Figure 6 As shown, Figure 6 This is a flowchart of steps S1-S4 of a stability control method for replacing primary pillars with artificial pillars in a phosphate mine goaf according to Embodiment 2 of the present invention. Based on the method described in Embodiment 1, the method of Embodiment 2 further includes the following steps:

[0219] Step S4: On-site industrial testing and full-process monitoring.

[0220] The solution determined in step S3 is applied to the field to establish a monitoring system covering the entire life cycle of replacement and service, and the monitoring data is fed back to correct the design parameters.

[0221] In step S4, artificial pillar casting and replacement of original pillars are carried out according to the scheme verified in step S3. Monitoring equipment is deployed to collect stress and displacement data throughout the entire cycle. The monitoring data is then checked in a closed loop with the design parameters of step S2 and the simulation results of step S3 to complete the stability control of the entire replacement process.

[0222] When performing artificial pillar casting and primary pillar replacement operations in step S4, an interval layout and skip-mining replacement construction strategy is adopted: the pillars are grouped into odd and even groups, the odd-numbered artificial pillars are constructed first and the corresponding primary pillars are removed after curing to the required standard, and then the even-numbered artificial pillars are constructed and the corresponding primary pillars are removed after curing to the required standard. Adjacent primary pillars are not removed at the same time.

[0223] In step S4, when deploying monitoring equipment, collecting data, and performing closed-loop verification with design parameters and simulation results, the following closed-loop monitoring operations are performed: stress gauges are installed inside the artificial pillar to collect stress data every 10 days; stress gauges and delamination meters are installed on the roof of the goaf to collect displacement data every 7 days; the safety factor and simulation accuracy are verified using the measured data. The specific process is as follows:

[0224] Step S4.1, Construction and stress monitoring of artificial pillars.

[0225] In this invention, artificial mine pillars are constructed using a layered on-site pouring method, with each layer not exceeding 500mm in height, and compacted using an immersion vibrator. After concrete pouring, a curing film is applied for moisture retention and curing for 28 days. When the concrete reaches 80% of its design strength, approximately 14 days later, adjacent primary mine pillars can be recycled. Primary mine pillars are dismantled one by one using mechanical cutting or static crushing methods.

[0226] Construction Sequence and Spatial Constraints: To ensure continuous support of the goaf roof during the replacement process, the construction of artificial pillars and the removal of primary pillars must follow the control strategy of interval arrangement and skip-mining replacement. Specifically: ① Divide the pillars into two groups according to odd and even numbers along the stope strike. First, construct all odd-numbered artificial pillars and cure them to the design strength, approximately 14 days; ② Once the odd-numbered artificial pillars have load-bearing capacity, the odd-numbered primary pillars can be removed; ③ Subsequently, construct the even-numbered artificial pillars and cure them to the design strength; ④ Finally, remove the even-numbered primary pillars. Throughout the entire process, no two adjacent primary pillars may be removed simultaneously, ensuring that the effective support spacing of the roof does not exceed twice the original support spacing at any given time.

[0227] An HCZ-2 oil-filled stress gauge is embedded inside an artificial rock pillar. The HCZ-2 oil-filled stress gauge is used in underground engineering projects such as mining and tunneling to measure stress changes in rock masses. It features a simple structure, direct stress reading, reliable performance, and convenient installation and use.

[0228] The stress gauge is installed at a designated depth inside the artificial pillar. After accurately placing the oil reservoir, monitoring personnel must remove the pressure cap from the gauge head, slowly loosen the screw at the gauge head, and use a manual pressure pump to steadily apply pressure to the preset value. After pressurizing, immediately tighten the screw at the gauge head and replace the pressure cap. Stress data is collected every 10 days during the subsequent monitoring period.

[0229] The stress monitoring results of the artificial pillars and roof of the North 2#778 mining area and the South 2#778 mining area are as follows: Table 15 shows the stress monitoring results of the goaf (partial representative data).

[0230] Table 15

[0231] 2025.9.10 4.83 4.97 3.52 4.02 Initial reading 2025.9.20 4.66↓ 4.81↓ 3.61 4.08 Neighboring area blasting disturbance 2025.9.30 5.14 4.97 3.74 4.22 2025.10.10 4.98↓ 4.85↓ 3.76 4.35 2025.10.20 5.35 4.99 — — 2025.10.30 5.56 5.01 — — 2025.11.10 5.45↓ 5.02 — — Artificial pillar maintenance completed 2025.11.20 5.76 5.01 — — 2025.11.30 5.76 4.89↓ — — 2025.12.10 5.75 5.02 — — 2025.12.20 5.76 5.01 — —

[0232] It should be noted that: ↓ indicates a slight decrease in stress value compared to the previous data collection period, with a maximum fluctuation range of approximately 0.25–0.50 MPa. This is mainly related to short-term construction factors such as disturbances from adjacent mining areas, equipment restarts, and backfilling operations. The overall data trend remains one of stable increase followed by a gradual leveling off.

[0233] like Figure 7 As shown, Figure 7 Figures show the stress variation curves of artificial pillars over time. Figure (a) shows the stress variation curve of the artificial pillar in the North 2#778 stope over time, and Figure (b) shows the stress variation curve of the artificial pillar in the South 2#778 stope over time. During the 120-day monitoring period, the stress in the artificial pillar in the North 2#778 stope showed an overall upward trend followed by a period of stabilization, but it was not a monotonically increasing trend. There were three identifiable small declines during this period, with the largest decline being approximately 0.35 MPa, which coincided with the timing of blasting operations and backfilling construction in adjacent stopes. The maximum stress value was 5.76 MPa, and the minimum stress value was 4.66 MPa. After entering November, the stress basically stabilized within a narrow range of 5.75–5.76 MPa. The maximum stress value of the artificial pillar in the South 2#778 stope was 3.76 MPa, and the minimum stress value was 3.52 MPa. After 30 days of monitoring, the stress basically leveled off. The monitoring data above indicate that the artificial pillars underwent a normal stress adjustment process during the mining disturbance and eventually entered a long-term stable bearing stage.

[0234] Step S4.2, stability monitoring of the goaf.

[0235] HCZ-2 oil pillow stress gauges and LBY-3 roof delamination instruments were installed on the roof of the goaf.

[0236] The roof stress monitoring results show that the maximum roof stress in the goaf of the North 2#778 mining area is 5.02 MPa, and the minimum is 4.81 MPa, with a small fluctuation range and a relatively stable curve trend. The maximum roof stress in the goaf of the South 2#778 mining area is 5.00 MPa, and the minimum is 4.02 MPa. Except for a few abnormal values, the monitoring results are relatively stable.

[0237] The LBY-3 roof delamination meter is a specialized device used in underground engineering fields such as mines and tunnels to monitor the displacement and delamination of roof strata. By strategically placing measuring points on the roof, this instrument provides real-time, all-weather monitoring of roof strata displacement changes, especially the occurrence and development of delamination. The LBY-3 roof delamination meter has a resolution of 0.01 mm, and its measuring range meets the requirements of this monitoring scenario.

[0238] The roof displacement monitoring method involves precise measurement and detailed recording of various data points of the roof in the goaf every 7 days. The roof displacement monitoring results are as follows: Table 16 shows the roof displacement monitoring results.

[0239] Table 16

[0240] 2025.9.1 North 2#778 goaf roof 0 0.2 0.2 2025.9.1 South 2# 778 goaf roof 0.1 0 0.1 2025.9.8 North 2#778 goaf roof 0 0.3 0.3 2025.9.8 South 2# 778 goaf roof 0 0.1 0.1 2025.9.15 North 2#778 goaf roof 0.2 0.1 0.3 2025.9.15 South 2# 778 goaf roof 0 0.3 0.3 2025.9.22 North 2#778 goaf roof 0.1 0.2 0.3 2025.9.22 South 2# 778 goaf roof 0.1 0.1 0.2 ... ... ... ... ... 2025.11.17 North 2#778 goaf roof 0.1 0.3 0.4 2025.11.17 South 2# 778 goaf roof 0.2 0.1 0.3

[0241] After 77 days of continuous monitoring, the amount of roof delamination in the goaf of the North 2#778 and South 2#778 sections remained within a small range of 0–0.5 mm, with the cumulative total delamination at each monitoring point not exceeding 0.8 mm, and no continuous unidirectional increasing trend was observed throughout the monitoring process. These monitoring data indicate that no significant cumulative delamination deformation occurred in the roof strata. This result is consistent with the predicted maximum roof displacement of 2.9 mm in the numerical simulation, further verifying the good stability of the roof under the support of artificial pillars.

[0242] Step S4.3, safety assessment.

[0243] During the 120-day continuous monitoring period of replacement and mining, all monitoring data met the quantitative safety standards in S3.2: the maximum roof settlement displacement was 2.9 mm, far less than 1 / 500 of the short-span of 15 m, i.e., 30 mm; the maximum measured stress of the artificial pillar was 5.76 MPa, while the maximum principal compressive stress of the artificial pillar in the numerical model was approximately 4.87 MPa. There was a deviation of approximately 18% between the simulation and the measurement, which falls within the reasonable error range between the numerical model and the actual rock mass mechanical response, typically <20%, mainly due to the simplification of model boundary conditions and the heterogeneity of the surrounding rock. The design used the field-measured maximum stress of 5.76 MPa as the control basis, corresponding to a short-term stress safety factor of 9.5 / 5.76≈1.65, meeting the short-term safety factor threshold of ≥1.5; the roof delamination fluctuated slightly within the range of 0–0.8 mm, without a continuous increasing trend. Based on this, it was determined that the stability of the stope in the early and middle stages of replacement was in a controllable and safe state.

[0244] Long-term stability verification: The initial monitoring data over the above 120 days verified that the artificial pillars possess good load-bearing capacity and stability in the short to medium term. Considering that the service life of phosphate mine goaf areas is typically several to several decades, concrete materials exhibit creep effects under long-term high stress, which may lead to strength degradation and deformation accumulation. Therefore, a creep reduction factor is introduced into the design. (This value is determined by referring to the provisions on the reduction of long-term strength of concrete under continuous load in the "Code for Design of Concrete Structures" GB 50010, and combining the empirical value of the creep coefficient of C25 concrete in a humid environment), and the effective strength of the artificial mine pillar is checked over a long period of time: .

[0245] Based on long-term strength verification, the stress safety factor of the artificial mine pillar is 8.08MPa / 5.76MPa≈1.40, which meets the long-term service safety factor threshold of 1.3 (this threshold refers to the safety factor provisions for permanent support structures in the "Technical Specification for Rock Anchors and Shotcrete Support Engineering" GB50086). This dual-threshold system, 1.5 for the short term and 1.3 for the long term, ensures the safety of the artificial mine pillar throughout its entire life cycle, including the construction and operation periods. It is recommended to conduct a comprehensive stress-displacement retest annually during the subsequent operation period to continuously verify long-term stability.

[0246] The above description is only a preferred embodiment of the present invention and is not intended to limit the present invention in any way. Any modifications, equivalent substitutions, improvements, etc., made within the spirit and principles of the present invention should be included within the protection scope of the present invention.

[0247] The above embodiments are merely illustrative examples and are not intended to limit the implementation. Those skilled in the art will recognize that other variations or modifications can be made based on the above description. It is neither necessary nor possible to exhaustively list all possible implementations. However, obvious variations or modifications derived therefrom are still within the scope of protection of this invention.

Claims

1. A method for stability control in replacing primary pillars with artificial pillars in phosphate mine goaf areas, characterized in that, The method includes the following steps: Step S1: On-site sampling and indoor rock mechanics tests are conducted on the phosphate mine to be replaced to obtain the physical and mechanical parameters of the surrounding rock and ore body; rock mass quality indicators are obtained through borehole exploration; a three-level grading system of rock mass geomechanical grading - improved rock mass geomechanical grading - corrected rock mass geomechanical grading is adopted to complete the fine grading of rock mass quality, forming a dual-track parameter of corrected rock mass geomechanical grading deterioration score and original rock mass geomechanical grading score; Step S2: Using the physical and mechanical parameters as the basis for calculating the stability of the primary pillar and designing the parameters of the artificial pillar, calculate the safety factor of the primary pillar and determine the initial stability state of the goaf; based on the contribution area method and the dual threshold safety system, quantitatively design the cross-sectional dimensions and material parameters of the artificial pillar. Step S3: Based on the dual-track parameters, the cross-sectional dimensions and material parameters of the quantitatively designed artificial pillar, a three-dimensional stope numerical model is constructed. The mechanical parameters of the three-dimensional stope numerical model are assigned by quantitative mapping through rock mass geomechanical classification, geological strength index, Hawke-Brown strength criterion, and Mohr-Coulomb strength criterion. The entire replacement process is simulated and the safety of the scheme is verified by quantitative criteria. If the conditions are not met, the process returns to step S2 for iterative optimization.

2. The stability control method for replacing primary pillars with artificial pillars in phosphate mine goaf areas according to claim 1, characterized in that, The method further includes: Step S4: Implement the artificial pillar casting and original pillar replacement operations according to the scheme verified in Step S3. Deploy monitoring equipment to collect stress and displacement data throughout the entire cycle. Perform closed-loop verification of the monitoring data with the design parameters in Step S2 and the simulation results in Step S3 to complete the stability control of the entire replacement process.

3. The stability control method for replacing primary pillars with artificial pillars in phosphate mine goaf areas according to claim 1, characterized in that, In step S1, the abnormal mechanical parameters obtained from the indoor rock mechanics test are reduced, and parameter range sensitivity analysis is carried out to avoid numerical model distortion caused by parameter anomalies.

4. The stability control method for replacing primary pillars with artificial pillars in phosphate mine goaf areas according to claim 1, characterized in that, In step S1, the three-level classification system of rock mass geomechanical classification - improved rock mass geomechanical classification - corrected rock mass geomechanical classification adopts a dual-track division of labor mechanism: First, the preliminary score of the rock mass geomechanical classification is calculated. Then, the improved rock mass geomechanical classification score is obtained after size effect correction. Finally, the improved rock mass geomechanical classification deterioration score is obtained by multiplying the five correction coefficients together. The original score of the rock mass geomechanical classification is used for subsequent parameter mapping, while the modified rock mass geomechanical classification deterioration score is only used for early warning of surrounding rock stability.

5. The stability control method for replacing primary pillars with artificial pillars in phosphate mine goaf areas according to claim 1, characterized in that, In step S2, the safety factor of the primary pillar is calculated using the Binyawski pillar strength formula, and the contribution area of ​​the pillar roof is divided using the area apportionment method. When the safety factors of all primary pillars are ≥1.5, the goaf is determined to meet the initial stability conditions for replacement.

6. The stability control method for replacing primary pillars with artificial pillars in phosphate mine goaf areas according to claim 1, characterized in that, In step S2, when quantitatively designing the cross-sectional dimensions of the artificial pillar, the minimum cross-sectional area of ​​the artificial pillar is quantitatively calculated using the following formula: ； in, This represents the minimum cross-sectional area of ​​an artificial pillar. A safety factor is designed for short-term use; This is a dynamic load correction factor; This refers to the vertical stress of the overlying rock strata. Contributes area to the top of the pillar; The effective strength of the artificial pillar in the field.

7. The stability control method for replacing primary pillars with artificial pillars in phosphate mine goaf areas according to claim 1, characterized in that, In step S2, the dual-threshold safety system is as follows: for the stage of artificial pillar replacement construction and mining impact, a safety factor of 1.5 is used for bearing stability verification; for the stage of stable operation of the goaf, a creep reduction coefficient is introduced to reduce the effective strength of the artificial pillar, and a safety factor of 1.3 is used for bearing stability verification.

8. The stability control method for replacing primary pillars with artificial pillars in phosphate mine goaf areas according to claim 1, characterized in that, In step S3, the quantitative mapping process is as follows: the original score of the rock mass geomechanical classification is converted into a geological strength index, the rock mass strength parameters are calculated by combining the Hawke-Brown strength criterion constant and the disturbance factor, and then converted into the Mohr-Coulomb strength criterion parameters by the tangent method. The artificial pillar is assigned values ​​using independent concrete mechanical parameters.

9. The stability control method for replacing primary pillars with artificial pillars in phosphate mine goaf areas according to claim 2, characterized in that, In step S4, when carrying out the artificial pillar pouring and primary pillar replacement operations, an interval layout and skip-mining replacement construction strategy is adopted: the pillars are grouped into odd and even groups, the odd-numbered artificial pillars are constructed first and the corresponding primary pillars are removed after curing to the standard, and then the even-numbered artificial pillars are constructed and the corresponding primary pillars are removed after curing to the standard. Adjacent primary pillars are not removed at the same time.

10. The stability control method for replacing primary pillars with artificial pillars in phosphate mine goaf areas according to claim 2, characterized in that, In step S4, when setting up monitoring equipment, collecting data, and verifying the closed-loop operation with design parameters and simulation results, the following closed-loop monitoring operation is performed: stress gauges are installed inside the artificial pillar to collect stress data every 10 days, and stress gauges and delamination instruments are installed on the roof of the goaf to collect displacement data every 7 days. The safety factor and simulation accuracy were verified using actual measured data.

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