Previous studies on mudstone caprock damage in underground gas storage (UGS) have largely focused on the macroscopic sealing failure, while the microstructural damage mechanism and the key controlling factors remain unclear. In particular, the microstructural damage has not been quantified systematically considering the stress levels and loading–unloading cycles. Therefore, this study developed a Particle Flow Code (PFC) numerical model calibrated by rock mechanics experiments to investigate the microstructural evolution in mudstone caprocks, taking the PG2 UGS in the Bohai Bay Basin as a case study. The results show that the maximum cycling stress, rather than the minimum cyclic stress or the loading-unloading cycles, exerts primary control over the microstructural damage in mudstone caprocks. Which can be reflected from the particle displacement velocity, the contact force chain strength and the microcrack number. An increase in maximum cyclic stress, a decrease in minimum cyclic stress, and a higher number of loading–unloading cycles will contribute to the propagation of microcracks and internal damage, ultimately leading to the failure of mudstone caprock sealing once the cycle count exceeds a critical threshold. Furthermore, a quantitative characterization model for microstructural damage was proposed, based on the relationship between crack density and strain. The model predicts that the maximum service life of mudstone caprocks is 160 cycles when the maximum cyclic stress ratio is 0.8 and the minimum cyclic stress ratio is 0.3. Additionally, the mudstone caprocks can withstand a maximum cyclic stress ratio of 0.96 when the loading–unloading cycles reaches 60. These findings provide a novel, quantitative basis for caprock stability analysis and operational parameter design in UGS.
Natural gas wells in gas-coal co-exploitation areas commonly penetrate thick unconsolidated aeolian sand, underlying bedrock, and longwall-mining-disturbed coal-bearing strata. Mining-induced subsurface deformation can transfer along the wellbore, threatening integrity at mechanical discontinuities. This study integrates theoretical analysis, 3D numerical modeling, and field-measured subsidence data to investigate how incompatible deformation and shear slip near the sand-bedrock interface translate into localized transverse shear loading. Results show that differential sand-bedrock movement induces active–passive earth pressure asymmetry, concentrating an additional transverse shear force of up to 71.6 kN near the critical interface, accompanied by a positive-to-negative reversal of lateral bending moment with peak values of 26.4 and − 24.8 kN m. Under sustained lateral loading, the cement sheath reaches its shear limit before the steel casing, and local cracking may compromise zonal isolation and increase gas leakage risk. Based on an upper-bound shear-capacity model considering in-situ strength reduction, the ultimate cross-sectional shear capacity of the cement sheath was estimated at 107 kN. Numerical simulations calibrated a linear relationship between surface subsidence and transverse shear force, with a transmission coefficient of 1.087 kN/mm and an R-squared value of 0.982, yielding a critical surface subsidence threshold of 98.4 mm. Under conservative limit-state assumptions with an empirical safety margin, combined with field-measured dip-direction subsidence functions, the recommended risk-control offset from the subsidence-basin centerline was 404 m. This risk zoning framework supports wellbore integrity assessment, gas leakage risk control, and mining-boundary optimization.
Early identification of potential roof collapse zones in coal mine roadways and implementation of effective prevention and control measures is of important practical significance for reducing the probability of roof fall accidents and preventing casualties. By investigating the sensitivity response relationship between drilling signals and the strength, thickness and assemblage of roof strata during the drilling process, this study reveals the attenuation law of vibration parameters and cutting parameters under drilled rock mass deterioration, and develops a method for identifying roof strata and potential roof collapse zones based on drilling parameters. The results show that drilling vibration and pressure responses are strongly controlled by specimen strength and integrity. From high- to low-strength specimens, the average vibration acceleration decreased from 1.367 to 0.865 g. The root mean square value decreased from 1.746 to 1.211 g, and the centroid amplitude decreased from 2578.294 to 534.828. The average pressure decreased from 12.954 to 7.387 kN, while the peak density decreased from 4.043 to 1.447 mm−1. Compared with intact specimens, fractured specimens showed clear attenuation in drilling vibration response, with attenuation rates of key vibration parameters ranging from 30.47 to 82.74
This study addresses the unclear distribution characteristics of advance abutment pressure and mining-induced stress evolution during deep coal mining. A theoretical model for advance abutment pressure was established by coupling the statistical damage constitutive relationship of coal-rock mass with roof displacement and horizontal stress evolution. The model incorporates the Weibull distribution, Mises strength criterion, strain equivalence principle, and an attenuation parameter to characterize the progressive damage behavior of coal-rock mass under mining disturbance. Parametric sensitivity analysis was conducted to evaluate the effects of Young’s modulus, Poisson’s ratio, mining height, and roof displacement parameters on the peak value and location of advance abutment pressure. A three-dimensional numerical model of Working Face 8301 in Chenmanzhuang Coal Mine was developed using Rhino and FLAC3D to investigate mining-induced stress evolution. Results indicate that the stress field evolves from the initial in-situ state to a pattern characterized by “goaf unloading–roof adjustment–stress concentration ahead of the face”. The advance abutment pressure exhibits a typical trend of rapid increase, peak development, and decay. The theoretical model predicts a peak stress of 44.80 MPa at 16.17 m ahead of the coal wall, while numerical simulations yield peak stresses of 42–48 MPa at distances of 5–17.5 m. The close agreement between theoretical and numerical results demonstrates that the proposed model can reasonably characterize advance abutment pressure distribution and mining-induced stress evolution under deep mining conditions, providing a theoretical basis for support design, mining safety assessment, and disaster prevention in deep coal mines.
The extraction of underground mineral resources can lead to stratum movement and deformation. Understanding these movement patterns is crucial for predicting surface deformation and mitigating geological disaster risks. However, the mechanisms behind surface horizontal displacement remain poorly understood, complicating the explanation of observed phenomena. This study uses numerical simulations to analyze horizontal displacement mechanisms caused by underground mining and develops a prediction model. The research identifies two causes of horizontal displacement: rotational displacement due to the inclination of the overburden at the goaf boundary, which correlates with surface inclination deformation, and translational displacement resulting from a “Z”-shaped movement pattern of the overburden during mining. This translational displacement concentrates in the central mining-influenced area and resembles ground subsidence. By integrating these two types of horizontal displacement, a new R T model is proposed. The R T model’s reliability is verified through 4 engineering cases, showing that the relative RMSE of the R T model is below 15
This research presents a newly developed mechanical model for analyzing surface movement dynamics and predicting displacement ranges resulting from sublevel caving operations in metalliferous deposits. A novel conceptualization of subsidence angle is introduced, accompanied by a calculation methodology derived from mechanical principles. The reliability of the new method was validated through the combined surface and underground monitoring approach implemented at the Dahongshan Iron Mine, along with numerical simulation calculations. The investigation examines how topographical variations, initial caved zone backfill elevation, and mining depth influence surface deformation patterns. Findings indicate that the reconceptualized subsidence angle—defined as the horizontal angle between the movement boundary extending from the deepest extraction level edge to the surface movement perimeter—creates maximum surface fracturing zones. Under flat terrain, subsidence zones manifest symmetrically on both sides of the undercut. Conversely, sloped landscapes produce asymmetrical subsidence patterns, with hanging wall subsidence angles exceeding footwall measurements. Nevertheless, due to the hanging wall topographic line intersecting the movement boundary at greater distances, hanging wall subsidence zones exhibit larger dimensions. While backfilling initial caved zones minimally affects tensile stress distribution areas, it diminishes stress magnitude, thereby inhibiting surface fracturing. Increasing extraction depths correlates with expanded tensile stress zones, though accompanied by decreased stress intensity. For combined extraction methodologies, the distribution of the principal stress on the surface, extending from the subsidence center towards the open-pit slope, can be sequentially divided into four zones: the compression stress decay zone, the tensile stress gradual increase zone, the tensile stress fluctuation zone, and the rapid increase zone of compression stress. Reduced slope angles effectively diminish overall tensile stress concentrations throughout the slope profile.
High-pressure hydraulic fracturing can enhance dense shale gas production, but it may reactivate faults through a decrease in frictional strength driven by both proximal and projected poroelastic stress, posing the risk of seismicity and even project shutdown. Hazard assessment is important for shale gas development. We explore the impact of poroelastic stresses on earthquake nucleation by deriving the critical characteristic slip length and nucleation length of a fault that rigorously incorporates both the magnitude and the rate of change in effective stress. We use this as an index to estimate the projected evolution from quiescent to seismic response of faults reactivated both during and after hydraulic fracturing operations. This work shows that increasing injection rate imposes dynamic loading on faults, reducing critical characteristic slip lengths and enhancing the potential to nucleate earthquakes. Conversely, reducing injection rates increases critical characteristic slip lengths and diminishes immediate seismic risk—allowing continued accumulation of tectonic stresses and the potential of larger delayed seismic events in the future. For sub-vertical faults, small events occur at shallow depths during injection with larger events on the deep fault post-shut-in. We interpret the seismicity observations from the Luxian shale gas field in the Sichuan Basin based on this assessment method. Pre-conditioning fracturing is suggested as a method to diminish the hazard of induced earthquakes, a risk that is also mitigated by greater well-to-fault separations and reduced injection rates. This work provides insights into fault slip behavior and seismic nucleation induced by hydraulic fracturing and guidance for its mitigation.
Water inrush from karst pipelines within coal seam roofs is a critical hazard constraining safe coal production in karst mining areas, yet its catastrophic evolution mechanism remains poorly understood. Taking the No. 21606 working face of Qinglong Coal Mine in Guizhou, China as the engineering prototype, this study independently developed fluid–solid coupling similar materials tailored for karst coal measure strata, and systematically explored the multi-field evolution patterns of the roof under mining disturbance through a combination of three-dimensional physical simulation and fluid–solid coupling numerical simulation. Physical simulation results demonstrate that the total development height of the caving zone and water-conducting fracture zone in overburden strata reaches 54.8 m. The roof stress exhibits a trend of initial increase, subsequent decrease and eventual stabilization, with a maximum roof displacement of 16.1 mm. Both pore water pressure and electrical conductivity remain stable at first, followed by a sharp rise and an abrupt drop, with peak values of 39.79 kPa and 87 ppm respectively. Numerical simulation results show that the roof plastic zone connects with the karst pipeline at a mining advance of 100 m. The maximum stress increment in strata near the coal seam reaches 2.65 MPa, and the maximum roof displacement is 52.80 mm. Pore water pressure peaks at 1.86 MPa at the location closest to the karst pipeline. This study identifies the ‘karst pipeline—lower-right fracture channel—working face’ as the dominant water inrush pathway, and reveals that the water inrush process follows a three-stage evolution law of stable stage—accumulation stage—water inrush stage. The findings can provide a theoretical basis and technical support for the prevention and control of roof water inrush in karst mining areas.
To reveal the coupled evolution of overburden fractures and pore structure during repeated slicing mining of an extra-thick coal seam, the 172,307 working face in Lingquan Coal Mine was investigated through physical similarity simulation and PFC2D numerical modeling. The staged failure of key strata, fracture propagation, strata pressure response, floor stress redistribution, and porosity evolution during upper- and lower-slice mining were systematically analyzed. The results show that the overburden response is strongly stage-dependent and closely related to key-stratum instability. During upper-slice mining, the fractured-zone height increased abruptly from approximately 50 m to 174 m following failure of the sub-key stratum at an advance distance of about 338 m. During lower-slice mining, failure of the primary key stratum at approximately 366 m triggered further high-level fracture development, and the fractured-zone height ultimately reached approximately 256 m. The average periodic weighting intervals were 30.4 m and 29.8 m for the upper and lower slices, respectively. Floor stress along the strike shows distinct zonation, including pressure relief near the open-off cut, a compacted bearing zone in the central goaf, and a strongly disturbed zone behind the working face. After lower-slice mining, the porosity ranged from 0.196 to 0.293 in the caving zone and from 0.142 to 0.295 in the fractured zone. Porosity evolution indicates that high-porosity zones migrate upward and concentrate near the open-off cut, stopping line, and upper part of the goaf, showing pronounced spatial reconstruction of the pore structure under repeated mining disturbance. These findings clarify the linkage among key-stratum failure, fracture evolution, stress redistribution, and porosity reconstruction in repeatedly disturbed overburden.
Deep coalbed methane resources hold significant potential and represent a critical area for achieving reserve growth and production increases of unconventional natural gas in China. To address the challenges in traditional sweet spot evaluation, such as difficulties in multi-parameter comprehensive assessment and heavy reliance on manual expertise, this paper establishes an intelligent sweet spot identification model that integrates a deep autoencoder with unsupervised clustering. Incorporating the coal structure index and ash content—critical factors directly influencing reservoir permeability and gas-bearing potential—along with key well-log parameters such as interval transit time and compensated neutron logs, nonlinear feature extraction and dimensionality reduction are performed using a deep autoencoder to capture complex interrelationships among logging curves. Furthermore, the K-Means clustering algorithm is integrated to automatically classify features within the encoded latent space, enabling intelligent delineation of sweet spot grades. Practical application demonstrates that the study area was automatically classified into Class I, Class II and Class III sweet spot zones. The optimal sweet spots exhibit reservoir characteristics of high interval transit time, high neutron porosity, high structural index, and low ash content. The prediction results from the intelligent evaluation method show an 84.35
The load-bearing and damage characteristics of coal seams vary under different geological conditions. To investigate the failure characteristics of coal bodies under hard roof conditions, this study is set against the backdrop of the Zhangji Coal Mine's 11,129 work face. By employing a combination of physical experiments, numerical analysis, theoretical calculations, and field validation, we explore the load-bearing capacity of coal bodies and their damage characteristics. The research findings indicate the following: (1) During the coal seam extraction process, the thick hard sandstone roof has a large span and exhibits minimal deformation, thus providing support for the overlying weak rock strata within a certain range. Following the implementation of pre-splitting blasting, the initial fracture step distance of the roof was reduced to 45 m, while the periodic fracture distance decreased to between 10 and 30 m. (2) The coal rock exhibits energy accumulation and dissipation at the moment of bearing stress or fracture. When the roof initially collapses, the peak stress within the coal body reaches 33.1 MPa, with a stress concentration coefficient of 1.7. The compressive displacement of the coal body within the stress concentration zone ranges from 2.8 to 12.1 cm. Energy accumulates within the coal body, reaching a maximum value of 124 kJ/m3, while the energy of the load-bearing layer above the goaf accumulates between 62.3 and 103 kJ/m3. (3) Based on the mechanical equations governing the load-bearing state of coal rock, a quantitative analysis was conducted on how four indicators—fracture step distance, distance from the rock layer to the work face, elastic modulus of the rock layer, and thickness of the rock layer—affect the coal body under dynamic and static load conditions, revealing the primary influencing factors. (4) Considering the geological conditions of the 11,129 work face, a pre-splitting blasting depressurization scheme was proposed using a fan-shaped hole grouping arrangement to reduce the fracture step distance of the hard rock layer. Upon the initial fracture of the roof, the support pressure in the lower section of the work face was consistently between 18.8 and 22.7 MPa. When the work face advanced through the pre-splitting blasting area (exceeding 700 m), the pressure in the middle support increased from 14.7 to 18.3 MPa to between 30.0 and 32.3 MPa, with a maximum recorded support pressure of 35.3 MPa. Field data indicate that the pre-splitting blasting scheme for the work face achieved the desired results. The study outcomes provide insight into the overall failure modes of coal rock bodies under the geological conditions of directly overlying hard roofs, thereby offering references for further investigation into the load-bearing characteristics of coal bodies and the calculation of crack propagation.
Rock-backfill composites, as key load-bearing structures in deep mining, exhibit mechanical behavior and failure mechanisms that are predominantly controlled by excavation-induced hole geometry and backfill conditions. In this work, a series of uniaxial compression tests were conducted on composite samples with various hole aspect ratios and backfill configurations using a coupled acoustic emission (AE) and digital image correlation (DIC) techniques. The results show that backfilling effectively mitigates local stress concentrations, enhances post-peak residual strength, and promotes the gradual evolution of fracture process. Increasing the hole aspect ratio weakens structural stability and leads to more uniform fracture morphology. AE analysis reveals a staged evolution of fracturing behavior and scale characteristics. The backfill structure shifts the energy release mode from abrupt to segmented, and increasing the backfill degree significantly suppresses high-frequency collapse-type events. Source mechanism analysis indicates that shear failure consistently accounts for a significant proportion, while a transition from tensile-dominated to collapse-dominated fracture occurs between tensile and collapse fractures. DIC observations clearly illustrate the regulatory effect of backfill on crack propagation paths, and displacement discontinuities at heterogeneous interfaces reflect the differential deformation between rock and backfill. These findings provide new insights into the geomechanical behavior of deep excavation-induced composite structures and offer practical guidance for optimized stope design, structural stability assessment, and multi-parameter early-warning strategies in underground engineering.
The shaft is a critical component in mine infrastructure, serving as the main conduit for access and material transport, and plays a pivotal role in maintaining operational stability. However, non-uniform pressure distributions pose a significant threat to shaft integrity, especially within the complex geomechanical context of deep multi-field coupling. To address this challenge, a depth-dependent 3D analytical model is proposed to evaluate the elastic response of deep shafts under non-uniform stress, with a specific focus on elliptical cross-section. The model establishes a 3D complex potential function framework for a depth-dependent stress field and derives the elastic stress–strain equation for elliptical shafts within depth-dependent in-situ stress field. Analytical solutions are obtained for stress and strain at typical burial depths of 300, 800, 1500, and 1800 m. A comparative analysis is conducted to investigate the influence of shaft geometry on its stability under non-uniform pressure. The elliptical axis ratio m and burial depth h significantly affect the stress and strain distributions around the shaft. At the unsupported shaft boundary, the radial stress is zero because of the traction-free boundary condition, whereas the hoop and vertical stresses vary with the boundary position, axis ratio, lateral pressure coefficients, and burial depth. The maximum hoop stress and the maximum radial, hoop, and vertical strains increase with increasing m, lateral pressure coefficient λ1, and burial depth h. Under the investigated stress conditions and model assumptions, elliptical shafts with smaller values of m exhibit lower stress and strain concentrations than the circular shaft, together with reduced compressive response and substantially lower tensile stress and strain. Furthermore, the analytical solutions show strong agreement with numerical simulation results. These findings provide valuable theoretical insights for the analysis, design and support of deep elliptical shafts.
To address the challenge of low rock-breaking efficiency in hard rock roadway, this paper proposes a method of pulse hydraulic fracturing to assist the rock breaking. First, fracture network by pulse hydraulic fracturing is formed in the hard rock. The rock is divided into rock blocks that can pass through the scraper conveyor, and damage occurs inside the rock block. Next, the cutting action of the roadheader’s cutter is employed to cause the rock blocks to easily break off. Additionally, by constructing observing borehole, the propagation range of the fracture network is controlled, so as to prevent the fracture network from causing great damage to the surrounding rock of the roadway. On-site trials were conducted in a mine's comprehensive excavation working face. Prior to pulse hydraulic fracturing, rock was ground into powder. After pulse hydraulic fracturing, rock could be cleaved along the fracture network, resulting in blocky rock fragments. The length and width of the rock blocks generally ranged between 10 and 90 cm. Dense microcracks appeared inside the rock block, with an average crack length of 3 mm, an average spacing of 22 mm, and an average opening of 81 μm. The number of fractures increased by approximately 46
To investigate the mechanical response of anchorage bearing structures in deep roadways under dynamic loading, a physical simulation test was conducted to analyze the evolution of the strain field under combined static and dynamic loads. A mechanical load-transfer model of the anchorage structure was established, and the key factors controlling instability under dynamic loading were identified. Results show significant stress relief in the roof and floor after excavation. During mining-induced stress loading, localized high-strain zones develop with a peak value of 12,000 με. Blasting disturbance leads to extensive stress release and vertical tensile cracks on the sidewall. Stress waves attenuate most when propagating from soft rock into hard rock. Increasing the pre-tightening force improves stability, shifts the peak tangential stress outward, and deepens the failure zone. Burial depth positively influences the bearing state but has limited effect on the failure location. Higher dynamic load intensity nonlinearly increases rock stress, raising the risk of large-scale dynamic failure. Field practice confirms that optimized anchor cables provide uniform stress distribution, ensuring safe mining operations.
The damage evolution mechanism of hollow sandstone under triaxial cyclic loading is crucial for evaluating the safety and stability of deep wells and deep tunnels. The strength, deformation, and damage evolution characteristics of hollow sandstone specimens (inner diameter of 15 mm and internal pressure of 0 MPa) were investigated by triaxial cyclic loading experiments and X-ray micro-CT scanning. Based on the experimental observations, a nonlinear parallel-bonded stress corrosion (N-PSC) model was developed in the particle flow code (PFC) to capture the damage behavior of hollow sandstone under cyclic loading. The fundamental principle of the N-PSC model is to reproduce cyclic damage by nonlinearly modifying the bond diameter. The results show that the peak strength of hollow sandstone under triaxial cyclic loading is lower than that obtained under conventional triaxial loading. The specimens exhibit three distinct stages of nonlinear damage development: initial damage, stable damage, and accelerated damage stages. Numerical specimens based on the N-PSC model can better reflect the macroscopic mechanical properties of hollow sandstone, including stress–strain curves, strength and deformation characteristics. The experimental energy analysis and numerical crack evolution jointly reveal the damage accumulation mechanism of hollow sandstone under cyclic loading. Microcracks initiate once the cyclic stress exceeds the crack damage threshold and rapidly propagate and coalesce in the post-peak stage to form a macroscopic fracture zone. Under triaxial cyclic loading, microcracks in hollow sandstone initiate from the inner hole wall and gradually extend towards the external wall. The specimens exhibit overall shear failure accompanied by hole spalling. At high confining pressure, stress concentrations at the specimen ends lead to severe end damage and localized failure. This research provides a theoretical explanation for the damage process and crack evolution mechanism of hollow rocks under cyclic loading.
Understanding the influence of anchorage parameters on bolt support effectiveness is crucial for optimizing support designs. Although prior studies have recognized the compensatory effect of bolts, the underlying mechanisms by which anchorage parameters modulate the support stress field and fracture behavior poorly understood. This study integrates laboratory experiments with discrete element method (DEM) simulations to investigate the spatiotemporal stress/energy evolution and fracture transition in bolt-reinforced coal ribs under compression. Results show that pre-tensioned bolts effectively mitigate coalburst by suppressing fracturing. Specifically, the crack initiation and damage stress thresholds increase logarithmically with pre-tension force but decrease with anchorage length. Concurrently, with the increase in pre-tension, the surface damage rate, the mass of far- and near-field coal burst-spalling debris, and the crack dimensions all exhibit a progressive decrease. Conversely, extending the anchorage length yields the opposite trends. Furthermore, DEM simulations reveal a distinct “two-compression-one-tension” support stress field coupled with an anisotropic force chain network, highlighting the asymmetric nature of load transmission and localized stress partitioning mechanisms. Crucially, both the magnitude and spatiotemporal evolution of the loading-induced support stress field are fundamentally dictated by the initial stress state, thereby providing new insights into the support design. Driven by the stress compensation-reinforcement mechanisms, the energy storage capacity and bonding strength of the coal ribs is enhanced, while the kinetic energy conversion ratio is significantly minimized. This elevates the critical threshold for coalburst initiation and diminishes the potential energy source for debris ejection, thereby successfully mitigating coalburst hazards.
Underground tunnels are inevitably exposed to dynamic disturbances during blasting excavation, yet the critical role of initial rock damage in controlling the long-term tunnel stability remains inadequately understood. In this study, a novel numerical framework is developed for simulating the sequential blasting excavation and subsequent creep process by embedding an enhanced dynamic viscoplastic damage model. The approach is then validated against an analytical dynamic excavation problem and field monitoring data from the Beishan Exploration Tunnel. The main contribution of this work is the quantitative assessment of how initial dynamic damage alters the time-dependent rock response via systematic simulations. In particular, our modeling results demonstrate that incorporating initial dynamic damage yields substantially more accurate predictions of time-dependent rock deformation, while neglecting the dynamic disturbance leads to a significant underestimation of creep development. The results further show that higher explosive pressures and faster unloading rates markedly intensify rock creep rates, whereas different in-situ stress conditions fundamentally alter the initial damage characteristics and ensuing creep evolution of the surrounding rock. Moreover, it is newly found that the pronounced local creep deformation can trigger non-negligible passive displacement in adjacent rock regions, highlighting the necessity of targeted reinforcement in highly disturbed zones to prevent cascading tunnel failure. The findings of our study offer practical insights for optimizing excavation design and support strategies of underground excavations in dynamically disturbed environments.
Injecting supercritical CO2 (scCO2) into deep saline aquifers is a promising strategy for large-scale CO2 emission mitigation. Although numerous studies have investigated scCO2 migration and displacement in porous sandstones, the dynamic displacement characteristics and the evolution of displacement efficiency under varying differential pressures remain poorly understood. Furthermore, few studies have distinguished displacement efficiency between dynamic displacement and equilibrium capillary state, limiting the accurate prediction of field-scale scCO2 storage efficiency. This study employs Nuclear Magnetic Resonance (NMR) and Magnetic Resonance Imaging (MRI) techniques to investigate the dynamic displacement of water by scCO2 in porous sandstone under pressure differences ranging from 0.1 to 22 MPa. Two experimental series were conducted, namely, (i) a single drainage displacement of water by scCO2 under various pressure differentials, and (ii) a three-stage process involving scCO2 injection, water imbibition, and scCO2 re-injection. By integrating displacement data with pore-size distribution characteristics derived from mercury intrusion porosimetry, the capillary pressure curve of the sample was determined. The results reveal that dynamic displacement yields higher water saturation than the static capillary equilibrium state by approximately 20
The shear behavior of rock joints is governed by the morphology and boundary conditions. However, most existing shear models cannot accurately capture the contributions of two-order asperities (waviness and unevenness) and non-stationary angles (i.e., the angle between the nominal plane and the shear direction, reflecting the uphill/downhill shear tendency). Here we propose a constitutive model incorporating the progressive degradation of three-dimensional asperities with explicit contribution of the non-stationary angle, based on the mobilized shear strength theory. The geometric parameters of two-order asperities are quantitatively separated, and the non-stationary angle is introduced into the constitutive relationship as a key joint property. By adopting a 3Dasperity degradation principle derived from the plastic tangential energy, an analytical shear model for rock joints is established. The proposed model is capable of modelling the entire shear process of joints. In the numerical implementation, the forward Euler method is employed for iterative solution, allowing simultaneous yielding of shear stress and dilation as a function of the shear displacement. Systematic laboratory direct shear tests were performed to investigate the mechanical responses of joints with distinct non-stationary angles under different normal stresses. The constitutive predictions agree well with the experimental results, demonstrating the validity and applicability of the proposed model. The involved parameters possess clear physical meanings and can be obtained from fundamental laboratory tests and 3D morphology. The proposed model may reliably represent the shear behavior of rough rock joints and evaluate pertinent rock-engineering stability.
