Surrounding Rock Deformation Control of Ore-Drawing Roadway Under Cyclic Ore-Drawing Disturbance

Abstract

Block caving is a cost-effective mining method that enables the highly efficient mining of thick and large ore bodies. During ore extraction in block caving operations, the ore-drawing roadways require especially high safety standards. However, the complex in situ stress conditions and cyclic loading from caved ore significantly deteriorate the stability of the surrounding rock. This makes rock mass control particularly challenging, such that it is crucial to study an effective method for maintaining the long-term stability of the roadways. This research proposes a comprehensive approach combining laboratory rock mechanics testing, numerical simulation, and field engineering validation to design effective support strategies for disturbance-affected roadways. Laboratory tests provide accurate mechanical parameters for the rock mass, the numerical simulations allow for the comprehensive analysis of deformation–failure mechanisms under disturbance conditions, and field validation ensures the reliability and practical applicability of the proposed support method. This study focuses on a −285 m ore-drawing roadway in the western section of the Yanqianshan Iron Mine. The in situ stress distribution was characterized through rock mechanics testing and acoustic emission monitoring. The propagation mechanisms of ore-drawing disturbance waves within the rock mass were analyzed, and numerical simulations revealed the deformation patterns and failure modes under dynamic disturbance, upon which the support scheme was designed. The results demonstrate that the designed bolt–mesh–shotcrete support scheme can effectively control surrounding rock deformation within 5 mm and resists the deformation induced by cyclic disturbances. This study provides valuable technical support for stability management in block caving mines with similar conditions.

1. Introduction

Block caving is a highly efficient, low-cost, large-scale mining method suitable for thick and massive ore bodies [1,2,3]. Its economic benefits can rival those of open-pit mining [4,5,6]. This method involves extensive undercutting to provide a free space for the caving of the overlying ore mass, utilizing the self-weight of the rock for ore transportation [7,8,9,10]. Within this system, the ore-drawing roadway serves as the primary channel for ore extraction. It is subjected to long-term high in situ stress and impact loads from the caved ore during drawing operations. Once damaged, these roadways are difficult to repair. In severe cases, this can lead to unrecoverable ore resources, endanger the safety of underground personnel, and cause significant economic losses to the mine. Therefore, ensuring the long-term stability of ore-drawing roadways is imperative for maintaining safe and efficient mine production.

Regarding ground stability issues in block caving mining areas, Xia et al. [11] investigated the damage and failure mechanism of roadways within the relatively weak “thin rock plate” at the base of block caving stopes. Shu et al. [12] found that the disturbances induced by caving processes during block caving mining can affect the geometric parameters and hydraulic properties of the surrounding rock mass. Furthermore, Xia et al. [13] identified that throughout the entire lifecycle of structures excavated using block caving, stability is influenced by multiple factors, maintenance proves challenging, and the risk of instability remains high. Qin et al. [14] found that during undercut advancement in block caving, compressive stress progressively concentrates ahead in the bottom structure, inducing shear failure risk, while stress significantly reduces in fractured zones during ore draw.

Concerning the control of surrounding rock deformation in roadways, Wang et al. [15] conducted research on hard roof control and optimized the design of its support parameters. Li et al. [16] employed a novel U-shaped steel set material for roadway support and validated its effectiveness using numerical simulation. Zeng et al. [17] studied the deformation and failure characteristics of deep soft rock roadways exhibiting poor surrounding rock properties and designed a support scheme based on a rational control strategy. Qian et al. [18] researched the primary influencing factors and failure mechanisms of argillaceous surrounding rock under high stress, leading to the development of innovative bolt materials and a corresponding support scheme. Ma et al. [19], addressing support failure caused by large deformations in deep soft rock, employed numerical simulation combined with field industrial trials to investigate the synergistic support effects of multiple support materials.

The aforementioned studies demonstrate that block caving mining significantly influences the mechanical properties of mine roadway surrounding rock. However, prevailing research predominantly relies on static assumptions, neglecting the impact of the distinctive continuous disturbance induced by the persistent caving and drawing of ore fragments—a hallmark of block caving—on roadway deformation. Existing dynamic investigations focus predominantly on the effects of blast vibrations or general mining disturbances on roadways [20,21,22,23,24], failing to encompass the high-frequency, long-term cyclic disturbances inherent to ore-drawing operations.

This paper takes a −285 m level ore-drawing roadway within the block caving zone in the western section of the Yanqianshan Iron Mine as the engineering case study, this research characterized surrounding rock mechanical parameters and in situ stress distribution patterns through laboratory rock mechanics testing. The disturbance characteristics induced by caved ore during drawing operations were analyzed. Numerical simulation was employed to investigate roadway deformation behavior and disturbance propagation patterns. Based on these analyses, an economical and safe support scheme was designed and implemented to control surrounding rock deformation. The field validation study confirmed the effectiveness of the scheme. These findings provide significant insights for the stability management of ore-drawing roadways subjected to continuous ore-drawing disturbances.

2. Geological Setting

The Yanqianshan Iron Mine is situated approximately 15 km from the urban center of Anshan City, Liaoning Province. It is a large-scale mining operation with over 50 years of history and represents China’s first iron mine to transition from open-pit to underground mining. The western section of the mine employs the block caving mining method. During the initial phase of this transition, as the block caving stope in the western orebody advances to greater depths, the ore-drawing roadways are subjected to the combined effects of complex in situ stresses, impact loads from caving rock masses, and repeated ore-drawing disturbances. This environment induces progressive cumulative fatigue damage in the rock mass, leading to time-dependent strength degradation and a significant reduction in the self-supporting capacity of the surrounding rock. Compounding this issue, the initial roadways were predominantly constructed using localized, temporary support schemes. Against this backdrop, research into the loading characteristics of the surrounding rock under the complex ground pressure conditions inherent to block caving, aiming to optimize the support system for ore-drawing roadways, has become an imperative for ensuring safe and efficient mining operations.

The western mining section of Yanqianshan Iron Mine is located west of the F19 fault as shown in Figure 1, striking NE with a dip angle of 70–85°. The structural features of the western section include a weakly weathered drill core exhibiting localized fracturing, fracture fillings predominantly composed of quartz, and joint surfaces characterized primarily by a rough, undulating morphology with subordinate smooth planar features. Rock Quality Designation (RQD) measurements indicate that magnetite quartzite in the western orebody ranges from 17.1% to 90.0%, with an average of 53.5%, with a mean joint spacing of 0.16 m, and an average joint density of 6.5 joints/m. Representative drill core photographs are shown in Figure 2. The rock mass in the mining area is generally classified as low basic quality, deteriorating to poor in severely argillized zones of the orebody. While the ore exhibits a relatively high strength, the overall stability of the orebody and its immediate roof and floor strata is moderate. However, in structurally complex areas with well-developed faults, intense jointing and fracturing result in fragmented rock masses, creating adverse conditions for mining operations.

3. Mechanical Properties of Surrounding Rock

3.1. Rock Mechanical Parameters

To characterize the mechanical properties of the surrounding rock, laboratory rock mechanics testing was conducted. Standard test specimens were prepared from 20 sets of drill cores obtained from the −285 m level of ore-drawing roadway in the western mining section of the Yanqianshan Iron Mine. Following the standard procedures of laboratory rock mass testing, uniaxial compressive strength tests, Brazilian splitting tests, and shear strength tests were performed on the specimens.

Through in situ sampling and laboratory rock mechanics testing, the rock mass mechanical parameters were characterized as follows: Yanqianshan −285 m ore rock specimens exhibited a uniaxial compressive strength of 116.08–181.68 MPa, elastic modulus of 20.10–97.32 GPa, tensile strength of 12.09–15.48 MPa, Poisson’s ratio of 0.15–0.20, with an average cohesion of 22.87 MPa and an average internal friction angle of 36.67°. The averaged mechanical parameters that are used in this study are listed in

Table 1. Average mechanical parameters of ore rock samples.

Elastic Modulus/GPa	Poisson’s Ratio	Compressive Strength/MPa	Cohesion/MPa	Internal Friction Angle/°	Tensile Strength/MPa
55.62	0.18	143	22.87	36.67	13.28

and as follows: compressive strength 143 MPa, elastic modulus 55.62 GPa, tensile strength 13.28 MPa, cohesion 22.87 MPa, internal friction angle 36.67°, and Poisson’s ratio 0.18.

3.2. In Situ Stress in Roadways

When rock is subjected to external forces, internal microcracks propagate and release elastic waves detectable by acoustic emission sensors. If the applied stress reaches or exceeds the historical maximum stress previously experienced by the rock, a sudden surge in energy release occurs, manifested as abrupt increases in acoustic emission counts and energy magnitudes. This phenomenon is termed the Kaiser effect. By analyzing acoustic emission signals, the historical maximum stress (i.e., in situ stress) can be determined [25]. Consequently, in situ stress conditions can be characterized through the laboratory acoustic emission testing of field-retrieved core samples.

Rock core sampling was conducted from the roadway floor at the −285 m level in the western mining section of the Yanqianshan Iron Mine. As illustrated in Figure 3, the oriented cores were first retrieved from the roadway floor in the field, and then small specimens (25 mm × 50 mm) in four orientations (vertical, and 0°, 45°, and 90° in the horizontal plane) were prepared through secondary coring in the laboratory. Utilizing the Kaiser effect of rock materials, uniaxial compression acoustic emission tests were performed on processed in situ stress testing specimens using an MTS testing machine coupled with an acoustic emission monitoring system. The system automatically recorded acoustic emission amplitude and absolute energy data. During cyclic loading–unloading and secondary loading sequences, cumulative acoustic emission counts and absolute energy values were statistically analyzed. The characteristic Kaiser point was identified by evaluating the Felicity ratio during secondary loading. By calculating stress values corresponding to Kaiser effect points from spatially oriented cores, the magnitude and orientation of the maximum horizontal principal stress at the measurement point were determined. Through the direct measurement of vertical principal stress and the indirect computation of two horizontal principal stresses, in situ stress values at varying depths were analyzed to establish regional stress field distribution patterns.

The experimental program employed an MTS 816 rock testing system integrated with a Physical Acoustics Corporation (PAC) acoustic emission monitoring system. Displacement-controlled loading was applied at rates of 0.05 mm/min and 0.1 mm/min during two-stage loading–unloading cycles. The acoustic emission system was configured with a 40 dB pre-amplifier gain and 45 dB noise threshold, as illustrated in Figure 4.

The stress components for different borehole orientations derived from acoustic emission testing are presented in Figure 5. The Kaiser point stress in the vertical direction corresponds to the vertical principal stress ${\sigma }_v$, with values detailed in

Table 2. Stress components in different borehole orientations.

Vertical Kaiser Stress (MPa)	Horizontal Kaiser Stress (MPa)
6.5	0°	45°	90°
	17.11	15.00	7.1

. For the three parameters in horizontal orientations, the maximum horizontal principal stress ${\sigma }_H$, the minimum horizontal principal stress ${\sigma }_h$, and the orientation of ${\sigma }_H$ were calculated using the planar principal stress equations (Equations (1)~(3)) based on elasticity theory. The results are summarized in

Table 3. In situ stress calculation results.

Vertical Kaiser Stress (MPa)	Max. Horizontal Stress σH (MPa)	Min. Horizontal Stress σh (MPa)	Principal Stress Direction θ (rad)	Angle (°)
6.5	17.39	6.82	0.16	9.35

.

$${\sigma }_H={\displaystyle \frac{1}{2}}\left({\sigma }_0+{\sigma }_{90}\right)+{\displaystyle \frac{1}{2cos2\theta }}\left({\sigma }_0-{\sigma }_{90}\right)$$

(1)

$${\sigma }_h={\displaystyle \frac{1}{2}}\left({\sigma }_0+{\sigma }_{90}\right)-{\displaystyle \frac{1}{2cos2\theta }}\left({\sigma }_0-{\sigma }_{90}\right)$$

(2)

$$tan2\theta ={\displaystyle \frac{2{\sigma }_{45}-{\sigma }_0-{\sigma }_{90}}{{\sigma }_0-{\sigma }_{90}}}$$

(3)

where ${\sigma }_0$, ${\sigma }_{45}$, and ${\sigma }_{90}$ represent the measured principal stress values at 0°, 45°, and 90° counterclockwise from true north, respectively; ${\sigma }_H$ and ${\sigma }_h$ denote the maximum and minimum horizontal principal stresses (compressive stress is positive, tensile stress is negative); and θ is the angle between the maximum horizontal principal stress and true north (positive in the counterclockwise direction from north).

3.3. Quality Classification of Surrounding Rock Mass

Based on the joint spacing, joint fissure conditions, Rock Quality Designation (RQD) values, groundwater conditions, and rock mechanical properties tests of the surrounding rock mass in the ore-drawing roadway at the western end of the Yanqianshan Iron Mine, the basic quality classification of the surrounding rock mass at the −285 m level ore-drawing roadway was determined to be Class III according to the Rock Mass Rating (RMR) classification system. The relevant parameters are shown in

Table 4. Calculation of basic quality classification of rock mass.

Parameter Values	RMR	Classification	Quality Description
Rock strength score = 12, RQD score = 8, joint spacing score = 8, joint condition score = 12, groundwater score = 7	47	III	Moderate

.

4. Numerical Simulation Study

4.1. Numerical Simulation Model

To comprehensively investigate the complex engineering and geological conditions of block caving mine roadways and their response under continuous ore-drawing disturbances, this study established a numerical model based on the designed roadway cross-section, support scheme, and mechanical parameters of the Class III surrounding rock obtained through field investigations and laboratory testing. An FLAC3D-based numerical simulation is used to analyze the displacement deformation of key load-bearing structures and plastic zone morphology development.

Taking the −285 m ore-drawing roadway west of the F19 fault as the research object, the cross-cut in the western section is arranged approximately perpendicular to the orebody strike, featuring a three-centered arch section with a 3.1 m straight wall height, 5.2 m clear width, and 1.7 m arch rise. Considering boundary effects, the model dimensions were set to 60 m × 32 m × 60 m (length × width × height). Local mesh refinement was implemented around the roadway. According to in situ stress distribution patterns, boundary stresses were applied: maximum principal stress 17.39 MPa, minimum horizontal principal stress 6.82 MPa, vertical stress 6.5 MPa, with a lateral pressure coefficient of 2.67. The model contained 145,920 elements and 154,473 nodes, adopting the Mohr–Coulomb constitutive model. Experimentally determined rock mass mechanical parameters were assigned, and the established finite element model is shown in Figure 6.

4.2. Stress Wave Induced by Ore-Drawing Disturbance

Ore-drawing disturbance is primarily caused by energy release during the impact of caved roof rocks on the ore pile in stopes above the roadway, which initiates localized alterations in the rock mass medium and triggers subsequent physical effects. Moreover, such disturbances alter the physical–mechanical properties of the surrounding rocks, particularly destabilizing geomechanically weak zones. These impacts necessitate research into effective support strategies to maintain roadway stability under cyclic loading conditions.

The orebody in the western section of the Yanqianshan Iron Mine is thick and relatively stable, with a near-vertical massive structure. At the −285 m level, the maximum caving height is 218 m and the average caving height is 189 m, indicating easy-to-moderate cavability. According to the ore-drawing conditions of the natural caving method at the western end of the Yanqianshan Iron Mine, numerical modeling was used to simulate the cyclic impact load caused by caved ore above ore-drawing roadways. The load waveform of dynamic disturbance waves in a rock mass can adopt a segment of harmonic waves in simulation [26,27]. Based on ore-drawing and extraction patterns, a cosine stress wave was selected to apply dynamic load to the model, as shown in Figure 7.

Stress waves propagate primarily as longitudinal waves in rock mass. The propagation velocity $C_p$ is calculated as follows:

$$C_p=\sqrt{{\displaystyle \frac{E}{\rho }}}$$

(4)

where E is the elastic modulus of the rock mass; $\rho$ is the rock mass density at 3380 kg/m³. The calculation yields a longitudinal wave velocity $C_p$ of 4056 m/s.

Since the positive Z-axis direction is upward in the model, while the disturbance stress from ore-drawing acts downward on the roadway, the expression for the drawing disturbance stress wave $P_t$ during simulation is as follows:

$$P_t={\displaystyle \frac{{\sigma }_d}{2}}\left(\cos \left(\omega \pi t_d\right)-1\right)$$

(5)

Field measurements indicate a peak drawing disturbance load ${\sigma }_d$ of 20 MPa and a dominant frequency ω of 100 Hz, with t_d being the disturbance duration. The Rayleigh damping model was selected to maintain accuracy during dynamic disturbance simulation. Based on geotechnical dynamic analysis experience, the damping ratio was set at 0.5%. To minimize stress wave reflection at model boundaries, free-field boundaries were applied to the static equilibrium model before dynamic analysis, absorbing incident waves to eliminate boundary effects. Considering the stope geometry above drawpoints, each experiment simulated 50 drawing disturbance cycles.

4.3. Failure Modes of Roadway Deformation

The field investigation and stability assessment of surrounding rock at the Yanqianshan Iron Mine reveal three primary failure modes in ore-drawing roadways—overall convergence, roof collapse, and rib spalling. Roof collapse occurs when fractured surrounding rock with developed joints undergoes local support instability, causing detachment under tensile stress. Rib spalling results from compressive–shear stress acting on fractured rock masses. Overall convergence manifests through the severe deformation of ribs and roof–floor interfaces or asymmetric pressure loading, leading to integral roadway contraction. Unsatisfactory long-term stability persists in unsupported or conventionally supported roadways, as shown in Figure 8, resulting in rational support strategies for ore-drawing roadway integrity maintenance.

The plastic zone from the numerical simulation of initial roadway deformation and failure is shown in Figure 9. Due to the high lateral pressure coefficient, the surrounding rock exhibits a marked rockburst tendency [28]. The plastic zone exhibits a butterfly-shaped distribution, indicating a progressive failure mode with a high strain concentration and extreme sensitivity to external disturbances, making it prone to malignant propagation. According to the simulated displacement results shown in Figure 10, shear failure dominates the surrounding rock damage. Tensile–shear failure occurs in shallow roof and floor areas. Maximum roof deformation exceeds 8 mm, floor deformation reaches 6.5 mm, and rib displacement reaches 4.3 mm.

After stress equilibrium, dynamic disturbance loads from ore-drawing were applied cyclically above the roadway for 50 cycles. To investigate the impact of drawing disturbance on roadway deformation, surrounding rock displacements at 10, 20, and 50 disturbance cycles were recorded for comparative analysis. The roof and rib displacements are shown in Figure 11 and Figure 12, respectively. During the initial drawing stages, the maximum roof displacement rapidly increased to 14.2 mm. At 20 cycles, it rose to 15.5 mm, stabilizing at 16 mm by 50 cycles. Rib displacement reached 6.1 mm after 10 cycles, gradually increasing to 6.8 mm with continued disturbance.

Displacement monitoring points were installed on the roof and ribs to observe the dynamic characteristics of roadway displacement. The relationships between displacement and the number of disturbance cycles at these monitoring points are plotted in Figure 13. The resultant displacement plots are shown in Figure 14. After 50 ore-drawing disturbance cycles, the maximum displacement rapidly increased during initial disturbance application before asymptotically approaching a stable value. Roof displacement increased from 8 mm to 16 mm, while rib displacement rose from 4.2 mm to 6.8 mm. These results demonstrate that early-stage caved ore disturbances significantly impact surrounding rock deformation, necessitating prompt roadway support implementation.

4.4. Roadway Support Scheme Design

To address the surrounding rock stability issues, four progressively reinforced support schemes were proposed based on ore-drawing roadway rock mass conditions and standard mining practices [17,29,30,31]:

① Plain shotcrete or bolt-only: 50 mm-thick shotcrete (≥C20 strength) or standalone bolts (φ 43 mm × 1800 mm; row and column spacing 1.0 m).

② Bolting with shotcrete or mesh: 100 mm-thick shotcrete (≥C20) supplemented by bolts (φ 43 mm × 1900 mm) or rock mesh (φ 6 mm reinforcement; 2 m × 1 m panels; 100 mm × 100 mm apertures).

③ Shotcrete–bolting–mesh support: GFRP bolts (φ 20 mm; tensile strength 1000 MPa; elastic modulus 50 GPa; 200 mm × 200 mm × 10 mm bearing plates). Rib bolts: 2.25 m depth; arch bolts: 3 m depth; spacing 0.8 m. Shotcrete: 100 mm thickness (C25 strength); mesh specifications match Scheme ②.

④ Combined support: Based on Scheme ③ with bolt spacing increased to 1.0 m and shotcrete thickened to 150 mm. Cable bolts (φ 20 mm × 5000 mm) installed at 800 mm × 1600 mm spacing. Light steel arches (0.75 m intervals) added in severely fractured zones.

4.5. Study on Support Simulation Results

The four support schemes were applied to established numerical models. Scheme ① used plain shotcrete, while Scheme ② employed bolting with shotcrete. The roof displacement simulation results are shown in Figure 15. As the support schemes were upgraded, maximum roof displacement decreased to 2.7 mm. The plastic zone nephograms in Figure 16 demonstrate reduced plastic zones at arch corners and ribs with enhanced support. The roadway failure mode transitioned from a progressive to a stabilized pattern. The plastic zone areas under each scheme are listed in

Table 5. Plastic zone area of roadway.

Support Scheme	Unsupported Scheme	Scheme ①	Scheme ②	Scheme ③	Scheme ④
Plastic zone areas (m2)	109.58	95.17	89.97	70.29	62.43

, confirming the effective control of plastic zone expansion and surrounding rock deformation.

As shown in Figure 17, the maximum displacement of the surrounding rock decreases with increasing support strength, while the impact of drawing disturbance weakens. The deformation fluctuations per drawing cycle stabilize progressively. The four schemes controlled maximum roof displacements at 13.21 mm, 11.43 mm, 4.22 mm, and 4.10 mm, respectively. To evaluate drawing disturbance effects, post-support static equilibrium displacement is defined as the initial value, and the ratio of maximum disturbed displacement to maximum static displacement is termed the disturbance displacement ratio (DDR), representing disturbance intensity. The specific values are listed in

Table 6. Displacement of surrounding rock.

Support Scheme	Unsupported Scheme	Scheme ①	Scheme ②	Scheme ③	Scheme ④
Plastic zone areas (m2)	109.58	95.17	89.97	70.29	62.43
Initial value (mm)	8.17	7.32	6.85	2.93	2.77
Max disturbed (mm)	15.92	13.21	11.43	4.22	4.10
DDR	1.95	1.80	1.67	1.44	1.48

.

4.6. Scheme Selection

A comparative analysis reveals that Schemes ① and ② reduced maximum roof subsidence by 10.4% and 16.2%, respectively, with plastic zone area reductions of 13.2% and 17.9%. These schemes only suppress shallow deformation with limited control effectiveness, exhibiting disturbance displacement ratios of 1.80 and 1.67 under cyclic loading, indicating significant displacement fluctuations and moderate disturbance resistance. Schemes ③ and ④ achieved substantial improvements, reducing maximum roof subsidence by 64.1% and 66.1% and plastic zone area by 35.9% and 43.0%, respectively. With disturbance displacement ratios values of 1.44 and 1.48, both maintained a maximum displacement below 5 mm during cyclic disturbance, demonstrating high stability and minimal disturbance impact.

Scheme ③ delivers control effectiveness comparable to Scheme ④ at a lower cost and steel consumption. Therefore, the shotcrete–bolting–mesh support (Scheme ③) is selected as the primary support method for ore-drawing roadways. Scheme ④’s combined support is reserved for severely fractured zones near weak interlayers or fault belts.

5. Engineering Application and Field Monitoring

To ensure the long-term stability of the −285 m ore-drawing roadways in the western section of the Yanqianshan Iron Mine, shotcrete–bolting–mesh support was immediately applied post excavation. The supported roadway is shown in Figure 18. To evaluate deformation control effectiveness, laser distance meters were installed at critical sections for real-time displacement monitoring and historical maximum value recording. The instrumentation layout is shown in Figure 19.

Field monitoring data recorded 90 days after support implementation revealed displacement differentials (maximum minus initial values) at each monitoring point, as shown in Figure 20.

The results demonstrate that the surrounding rock deformation in supported ore-drawing roadways was controlled within 5 mm, with displacements below 2 mm in 3# roadway. The high-strength GFRP bolts effectively restrained butterfly-shaped plastic zone expansion. Dense mesh and C25 shotcrete synergistically attenuate drawing disturbance stress waves, ensuring favorable surrounding rock safety and long-term ore-drawing roadway stability.

6. Conclusions

To address the long-term stability requirements of ore-drawing roadways in block caving method applications, this study focused on the deformation control of roadways surrounding rock under ore-drawing disturbance conditions. This research proposes a comprehensive approach combining laboratory rock mechanics testing, numerical simulation, and field engineering validation to design effective support strategies for disturbance-affected roadways. The main conclusions are as follows:

(1) Random core sampling from ore-drawing roadways in the western section of the Yanqianshan Iron Mine was conducted, followed by comprehensive laboratory rock mechanical tests. The surrounding rock was classified as Class III according to the RMR classification system. Based on rock mechanics experiments and acoustic emission Kaiser effect testing, the in situ stress state and mechanical parameters of the surrounding rock were determined. Due to the elevated lateral pressure coefficient of 2.67, the surrounding rock exhibits a marked rockburst tendency.

(2) For the block caving operation, the key parameters influencing drawing disturbance on ore-drawing roadways and waveform characteristics were analyzed. Critical simulation parameters for drawing disturbance were determined. Roadway response was simulated under disturbance, with subsequent displacement monitored. The disturbance displacement ratio (DDR) was proposed to quantify drawing disturbance impact on roadway deformation.

(3) Numerical simulation analyzed the deformation patterns of key bearing structures in ore-drawing roadways by monitoring surrounding rock displacement and plastic zone morphology. Significant displacement occurred at the roof and ribs, while butterfly-shaped plastic zones exhibited high susceptibility to disturbance-induced expansion. Four support schemes were designed for ore-drawing roadways, with bolting–mesh–shotcrete selected through simulation to control deformation and failure.

(4) The implementation of the support scheme at −285 m ore-drawing roadways in Yanqianshan’s block caving operation was validated through multi-point displacement monitoring. Field measurements confirmed post-support displacements below 5 mm, demonstrating effective ground control that meets long-term safety and stability requirements for mine production.

Although this study provides a comprehensive solution for supporting ore-drawing roadways, some limitations and future research directions remain. Firstly, the dynamic disturbance load in the simulation was simplified as a cosine wave; future work could incorporate more realistic load waveforms measured from field monitoring. Secondly, the fatigue effects on support materials under long-term cyclic loading were not thoroughly investigated; future studies should focus on the durability and lifespan of support systems. Lastly, the research could be extended to develop a real-time monitoring and early warning system based on the findings of this study to further enhance the safety of mining operations.