In engineering practice, surrounding rock is commonly dissected by joints and thus becomes discontinuous, which can readily trigger instability and failure. Prestressed rock bolts provide effective reinforcement for jointed surrounding rock; however, current design still relies largely on experience and engineering analogy. Therefore, a quantitative evaluation of bolting effectiveness is necessary to guide engineering practice. By combining theoretical analysis and numerical simulation, the prestress-induced stress compensation of rock bolts is quantified. The results show that: (1) for a single bolt, the peak vertical stress compensation is approximately 20 kPa in intact rock and decreases to about 16 kPa when a joint is present; (2) for a three-bolt system, stress superposition forms a reinforced compressive zone, and the peak vertical stress compensation increases to about 35 kPa in intact rock and 27 kPa in jointed rock; (3) on the joint surface, the normal and tangential stress compensations increase from about 13 kPa and 3 kPa (single-bolt case) to about 25 kPa and 7 kPa (three-bolt case), respectively; (4) at the tunnel scale with systematic bolting, prestressing increases the near-boundary radial stress by approximately 0.2-0.3 MPa and improves global stability; (5) although the absolute magnitude of direct stress compensation is modest compared with rock strength, direct shear simulations indicate a clear strengthening mechanism: under a 100 kN prestress level, the equivalent cohesion increases by about 13.5% and the internal friction angle increases by about 40.3%, leading to a marked improvement in the overall stability of jointed surrounding rock.

To reveal the stress compensation effect of prestressed rock bolts in the surrounding rock with structural planes, model tests under non-support and high prestress support conditions were carried out. The deformation and failure laws, failure mechanisms, and relative movement states of surrounding rock blocks were compared and analyzed. The function mechanism of prestressed rock bolts in jointed surrounding rock was obtained, considering the stress compensation effects of different rock bolt lengths and prestress values on jointed surrounding rock. The results show the following: (1) Prestressed rock bolts formed a Supporting structure with the tunnel surrounding rock, thereby increasing the ultimate load of the model from 0.42 to 0.6 MPa, with a 42.8

During the construction process of blocky rock mass tunnels, the phenomenon of falling arch blocks frequently occurs. To truly reflect the deformation characteristics of the surrounding rock, an interlocking structure block model method considering the distribution characteristics of the surrounding rock structural planes was proposed, and the rationality of the structured block model was validated through numerical calculations. First, by conducting an interlocking structured block model test, the deformation and failure characteristics of the tunnel section during the excavation process were analyzed, and the displacement of the surrounding rock clearly exhibited asymmetrical deformation. Owing to the interlocking effect of nonpenetrating blocks, the torsion and subsequent gyration of the blocks limited the range of surface damage to the surrounding rock, ultimately resulting in shear fracture during the overload test. Second, compared with the phenomenon of small deformation of the surrounding rock in the continuous medium model, whose bearing capacity is 9 times that of geostress, the nonpenetrating structural plane inlaying structure block model formed a pressure arch structure with a height of one quarter of the tunnel diameter, resulting in a bearing capacity that is twice as high as the geostress, which is closer to the actual engineering situation. In addition, the prestressed rock bolt support technology effectively controlled the deformation of the surrounding rock of the block test model, and the deformation of the surrounding rock was controlled at 6.5 mm through field testing. The block test model accurately reflects the variation characteristics of jointed surrounding rock and provides a reference for tunnel support design.

Abstract Rock blocks formed by joint cutting at the tunnel arch crown are prone to collapse and instability. Considering the in situ joint distribution characteristics, this study investigates the controlling effect of prestressed rock bolts on overhanging rock blocks at the tunnel arch crown and the influence of rock bolt parameters. The effectiveness of prestressed rock bolts in controlling overhanging rock blocks is verified through a model test. The results indicate that: (1) Prestressed rock bolts can effectively restrain block sliding and provide radial confinement. Compared with ordinary rock bolt support, prestressed rock bolt support increases the joint normal stress around the overhanging rock block and significantly reduces the difference between the major and minor principal stresses, thereby increasing the strength reserve of the rock mass. (2) Increasing the rock bolt prestress and length and appropriately reducing the rock bolt spacing can reduce the settlement of the overhanging rock block, improve the stress state of the surrounding rock, and enhance the normal constraint across the joint planes. Once the rock bolt prestress reaches 100 kN, the rock bolt length reaches 4.5 m, and the rock bolt spacing is reduced to 1.3 m, further changes in the rock bolt parameters provide only limited additional improvements. A relatively reasonable prestressed rock bolt support scheme consists of a rock bolt prestress of 100 kN, a rock bolt length of 4.5 m, and a rock bolt spacing of 1.3 m. (3) The model test shows that the prestressed rock bolts cross the joint planes and provide effective anchorage, thereby restraining the settlement of the overhanging rock block and preventing its instability. (4) Field monitoring results show that the tunnel crown settlement does not exceed 4 mm, the rock bolt axial force stabilizes within the range of 95–110 kN, and the overall stability of the tunnel is maintained. The research results provide an effective support method for tunnels with similar joint distributions.

To clarify the supporting effect and influencing factors of energy-absorbing rockbolts in soft rock roadways with large deformation, by considering four factors including rock deformation, plastic zone, rockbolt force, and stress of the surrounding rock, comparative analysis of no-rockbolt, conventional rockbolt, and energy-absorbing rockbolt schemes is conducted. The effect of the energy-absorbing rockbolt is analyzed based on a self-developed numerical simulation program, and a study is conducted on the influence of five factors such as the energy-absorption starting axial force, ultimate yielding distance on the supporting effect. The results show that: 1) Compared to conventional rockbolts, the energy-absorbing rockbolts maintain a intact support system and continuously providing support resistance within 75 d of calculation. 2) Energy-absorbing rockbolts significantly increase the maximum and minimum principal stresses of the roadway. The increase in maximum principal stress significantly enhances the range of the bearing arch in the surrounding rock. Energy-absorbing rockbolts have a high capacity to compensate for radial stresses unloaded. 3) The greater the energy-absorption starting axial force, the more significant the compensation effect of the rockbolts on the radial stress. Appropriately increasing the ultimate yielding distance and rockbolt length can effectively prevent rockbolt failure and control the area of plastic zone.

A mechanical model of the stress distribution of the protective coal pillar was constructed to explore the reasonable width of a protective coal pillar in an isolated working face and further reduce the impact of hazardous gases on the mining face, taking the mining of the isolated working face under the goaf as the engineering context. The width of the protective coal pillar was positively correlated with the height of the key layer and the stress of the accessories and negatively correlated with the ultimate bearing strength and caving angle of the coal pillar. The stress distribution of the surrounding rock under different coal pillar widths is analyzed by a numerical model. Field tests revealed that the deformation of the surrounding rock was effectively controlled when the width of the protective coal pillar was 23 m. The proportion of the elastic core area inside the coal pillar is 46.8%, and the overflow of hazardous gas is essentially zero within 30 days. This research provides a reference for the retention of protective coal pillars and the control of hazardous gas under similar engineering geological conditions.

In order to explore the deformation and failure characteristics of jointed surrounding rock in ultra shallow buried tunnel, a interlook test block was constructed, which conforms to the distribution characteristics of joints. The rationality of the block model is verified by the mechanical properties of rock mass. In the process of tunnel excavation, the trapezoidal settlement area of the surrounding rock in the arch is transformed into a rectangular settlement area, and the blocks in the left area of the tunnel arch are relatively sliding due to the joint inclination. It is further clarified that the control effect of prestressed anchor on rock mass in jointed surrounding rock is significantly higher than that of non prestressed anchor. The prestressed rock bolt support effectively controlled the deformation of surrounding rock, and reduced the area of plastic zone by 54.47 %. The influence of the length and spacing of prestressed rock bolts on the deformation of surrounding rock is further analyzed. The area and displacement of tensile stress zone and plastic zone decrease with the decrease of rock bolt spacing. When the spacing is less than 1.2 m, the area of tensile stress zone suddenly decreases, and the plastic zone decreases by 38.8 %. With the increase of bolt length, the tensile stress zone, plastic zone and displacement decrease. When the rock bolt length is greater than 3.5 m, the tensile stress zone suddenly decreases, the plastic zone decreases by 42.3 %, and the displacement decreases by 73.6 %. Applying the support scheme to the construction site, the deformation of jointed surrounding rock is finally stabilized at 4 mm. This work provides a reference for the support design of jointed surrounding rock in ultrashallow buried tunnels.

The existence of jointed rock masses in tunnels is a key problem that leads to the deformation and instability of the surrounding rock. Based on the engineering background of the Qingdao Metro, a physical model and a numerical model are constructed to analyze the mechanism of deformation and instability of the surrounding rock of a jointed tunnel using model blocks that conform to the characteristics of joint distribution. The results show that during the whole test process, the cracks between the blocks in the tunnel arch are in the "development - stability - closure - penetration" state. Finally, the block falls in the left area of the tunnel, and the corresponding anchor axial force transits from the linear growth stage to the steady growth stage. After the axial force of the rock bolt suddenly decreases in the loading stage, the bolt enters the stability stage, and finally, S-shaped failure occurs when the surrounding rock is unstable. Under the action of a prestressed rock bolt, the normal stress and radial stress of the rock mass increase, and the difference between the major and minor principal stresses decreases significantly. The radial stress on the surrounding rock surface of the arch is approximately 8 % greater than that in the traditional rock bolt scheme, which improves the strength reserve of the rock mass. In the field test, the prestressed rock bolt can control the deformation of the surrounding rock in the arch within 2 similar to 5 mm, which provides a reference for the control of the surrounding rock in jointed tunnels.

In response to the problem of severe deformation and failure of the surrounding rock of the mining tunnels in inclined strata, a tensile-shear failure numerical model of rock bolts was adopted to analyze the rock bolts failure behavior and the surrounding rock mechanical response, and explore the influence of different factors on rock bolts failure. The results indicate that: (1) As angle between the rock bolt and the interface decreases, the shear force of increases, and the rock bolt is easier to fracture. The maximum displacement is 250.2 mm of fractured position. A new evaluation index of shear axis ratio was proposed. The larger the shear axis ratio, the rock bolt is easier to fracture. The shear axis ratio at the interface has increased by 95 % compared to the non interface. (2) The plastic zone and the surrounding rock deformation exhibit the same inclination trend as the strata. In final state, the maximum deformation is 297.7 mm at the bottom and top position perpendicular to the rock bolts at the interface. At the moment of rock bolts fracture, the interface displacement of surrounding rock immediately increased by about 10 mm. And the stress in the surrounding rock decreased about 1 MPa. The weakening effect of surrounding rock stress shifts with the change of rock bolt fracture position. The stress at the arch waist decreases by 44 %, the stress at the arch crown decreases by 25 %. When the fracture of the rock bolts at the arch crown and arch waist ends, the stress at the side wall begin to decrease by 40 %. (3) As the inclination angle increases, the deformation of the surrounding rock and the number of rock bolt failures increase. The range of the plastic zone in the surrounding rock gradually increases, and the degree of rock bolts asymmetric fracture more aggravated. As the diameter of the rock bolt increases, the fractured number of the rock bolts and range of plastic zone decreases.

Under the action of tectonic stress, the stress release during the construction of shallow buried massive rock tunnel often leads to the change of surrounding rock displacement mode. Based on the phenomenon of upward movement of the ground surface during the tunnel excavation of Qingdao metro, combined with the distribution characteristics of surrounding rock structural plane, a new block test model is constructed, and it is clear that the displacement mode of surrounding rock of shallow buried tunnel deflects upward under the action of tectonic stress, but there is a 1.2-3.5 cm collapse area on the surface of surrounding rock of the tunnel. The numerical simulation shows that with the increase of tectonic stress level, the displacement mode of surrounding rock of shallow buried tunnel develops from the settlement type at 0 MPa to the rise and fall type at 3 MPa, and finally becomes the rise type when the tectonic stress exceeds 7 MPa. Through theoretical analysis, it is found that the strength of surrounding rock bearing structure under the action of prestressed bolt support is mainly related to the diameter, length and prestress value of the bolt, as well as the spacing between rows of bolts. Based on the fact that the shallow buried tunnel of Qingdao Metro is in the displacement mode of both rise and fall, increasing the bolt length is conducive to enhancing the synergy between the rock surface and the deep. After field verification, the surrounding rock deformation of the tunnel is controlled within the range of 2-5 mm. The research results provide a basis for the selection and control of surrounding rock support parameters under similar geological conditions.

The lattice girder (LG) and shotcrete work together to provide support in tunnel engineering, being influenced by the hardening process of concrete and the excavation process. To elucidate the dynamic changes in the mechanical behavior of the LG during this process, the present study conducted indoor experiments using a self-developed loading system suitable for early age flow-plastic state shotcrete LG. Through the secondary development of the ABAQUS concrete hardening process constitutive model, an analysis of parameters such as curvature and load values was conducted to understand the evolution of the load patterns on the LG. Subsequently, relying on the background of the construction of an underground excavation station in Qingdao Metro, refined numerical simulations of the construction process were carried out. Combining the analysis with measured data, the following conclusions were drawn: 1. In the first three days after concrete casting, the LG specimen experiences relatively low stress, and after seven days, the LG specimen and concrete form a cohesive synergy; 2. The initial load caused the deterioration of the supporting performance of the unhardened grid concrete specimens, and the bearing capacity of the test group was reduced by 25% compared with the control group; 3. In the early stages of tunnel excavation and support, due to the low hardening process of shotcrete, the ability to transfer the surrounding rock load to the LG is limited, and the layout of the LG has no significant impact on the stability of the surrounding rock; 4. The collaborative support effect between the rebar bent frame (RBF) and concrete is greater than that of the LG, and it coordinates better with prestressed anchor bolts, presenting advantages in hard rock tunnels.

At present, there are a variety of concrete-filled steel tube (CFST) supporting arches with different cross-sections, but a method for selecting a section that fits the site conditions is still lacking. In this paper, laboratory tests and numerical tests were carried out, and three mechanical conditions, the local compression-bending loads, the simplified ideal global loads and the simulated arch-rock contact loads, were investigated for the three most commonly used cross-section types, namely circular, square and D-shaped cross-sections. For the local compression-bending capacity, the axial bearing capacities of D-shaped cross-section specimens are 1.45% and 4.58% higher than those of the circular and square cross-section specimens, respectively; the bending bearing capacities of the square cross-section specimens are 11.19% and 22.13%/21.27% (D-shaped specimens under reverse eccentric compression) higher than those of the circular and D-shaped cross-section specimens, respectively. For the simplified ideal global loads, the straight-leg arches with square, circular and D-shaped cross-sections have bearing capacities from high to low, while the order of the bearing capacities is circular, square and D-shaped for circle arches; the difference values are not more than 10%. For the simulated arch-rock contact loads, the bearing capacities of the D-shaped type arches are 12.42% and 5.89% higher than those of the circular and square arches, respectively, under the full contact condition; the stress distribution of the surrounding rock is irregular, the stress concentration phenomenon appears near the cavity under the non-full contact condition, and the square cross-section arches have better bearing capacity behaviour with cavity widening. Finally, suggestions for the cross-section selection of the CFST supporting arch are proposed according to the quantitative test data and combined with the processing technique, production cost, construction technique and connecting joint property.

To realize the numerical analysis of rock bolt bearing-yield-fracture under tension-shear coupling and to provide an effective means for supporting the design of roadways or tunnels, a new numerical model was developed to analyse the tension-shear coupling fracture behaviour of rock bolts in engineering rock masses. Based on the study of the mechanical behaviour of rock bolts under tensile-shear coupling load, the mechanical criterion and constitutive model of the yield and fracture of rock bolts were proposed, and the secondary development of the FLAC3D platform was realized. At the same time, self-developed equipment was used to conduct the rock bolt tensile shear test. By comparing the numerical simulation results with the experimental results, we determine that the maximum difference in the shear displacement of the rock bolt does not exceed 2 mm during the yielding and breaking of the rock bolt, and the percent difference was approximately 3% and 10%, respectively. In the field of engineering, this result is generally considered acceptable. Using the numerical model we proposed, a numerical simulation experiment was conducted on the rock bolt support design of a coal mining roadway in engineering practice. The results show that under the original support scheme, a large amount of rock bolt tensile-shear failure occurs at the coal-rock interface, which makes the original support scheme unable to meet the engineering requirements with tensile-shear coupling failure tendency. In this regard, we adopted the support method of grouting anchorage at the interface of coal and rock. The results show that the deformation of the roadway surrounding rock has been effectively controlled.

The filling mining method is important in realizing the green mining of mineral resources. Aiming at the problems of land resource occupation, environmental pollution, and rational utilization of coal-based solid wastes such as coal gangue, fly ash, and desulfurization gypsum, a new paste filling material was developed with coal gangue, fly ash, and desulfurization gypsum as raw materials. The microstructure of the raw materials was analyzed by XRD and SEM. Combined with the Box-Behnken experimental design, the effect of each component on the fluidity of the filling slurry was analyzed through the response surface analysis. The significance of each component on its bleeding and fluidity was determined, and the optimal ratio of the filling slurry was obtained. Experimental results show that the microcosmic morphology of coal gangue, desulfurization gypsum, and gasification slag presents an irregular block and rough particle surface; the microcosmic morphology of fly ash and bottom slag presents first out spherical or quasi spherical particles. Moreover, obvious sintering traces exist on the surface of the bottom slag. The main crystal mineral of coal gangue and fly ash is SiO2, the desulfurization gypsum is composed of Ca(SO4) (H2O) and Ca(CO3) crystal minerals, the gasification slag is composed of carbon and nitrogen compounds, and the main crystal mineral components in the bottom slag sample are SiO2 and AlxSiyOz compounds. The order of significance of each key factor on slurry fluidity is as follows: C (desulfurization gypsum) > D (gasification slag and bottom slag 1:1) > A (coal gangue) > B (fly ash). The order of the significance of each key factor on slurry bleeding is as follows: B (fly ash) > C (desulfurization gypsum) > D (gasification slag and bottom slag 1:1) > A (coal gangue). Considering the material preparation, field application, and other conditions, the mass percentage of each factor content of the new paste filling material is as follows: 49.5% coal gangue, 8.3% fly ash, 4.1% desulfurization gypsum, 6.2% gasification slag, and 6.2% bottom slag.

In elastic mechanics, the complex function method proposed by Muskhelishvili can be used to solve the stress distribution of rock surrounding a roadway with an irregular cross-section. However, the solution of the conformal mapping function from the exterior of the roadway to the interior of the unit circle is a prerequisite, but it is difficult to obtain. In this study, based on the Riemann mapping theorem and the boundary correspondence principle, the conformal mapping function was approximated using a Laurent series with finite terms. Assuming that the polar angle of the two corresponding points on the image of the conformal mapping function and the boundary of the roadway cross-section are equal, an iterative algorithm for calculating the conformal mapping function is proposed using the least squares method, and the code was programmed by-using Python. Using the proposed algorithm, two conformal mapping functions were solved for roadways with irregular cross-sections in practical engineering projects, while the parameters and the error were examined. Simultaneously, the performance was analyzed statistically for curved and broken line boundaries. The analysis results has shown that the errors were concentrated on the corners of the roadway. In addition, it was observed that an excessively large number of sample points would not improve the accuracy, but will only extend the algorithm convergence time. When the series includes more terms, the accuracy will be higher and the convergence speed will be slower. However, statistical analysis has shown that the algorithm was converged rapidly, with convergence time of less than 10 ms. Finally, the effectiveness of the algorithm was verified by solving the stress distributions of the rock surrounding two roadways. The algorithm might be used to solve the conformal mapping function from the exterior of a roadway with an irregular cross-section to the interior of the unit circle in coal mines.

In coal mining roadway support design, the working resistance of the rock bolt is the key factor affecting its maximum support load. Effective improvement of the working resistance is of great significance to roadway support. Based on the rock bolt's tensile characteristics and the mining roadway surrounding rock deformation, a mechanical model for calculating the working resistance of the rock bolt was established and solved. Taking the mining roadway of the 17102 (3) working face at the Panji No. 3 Coal Mine of China as a research site, with a quadrilateral section roadway, the influence of pretension and anchorage length on the working resistance of high-strength and ordinary rock bolts in the middle and corner of the roadway is studied. The results show that when the bolt is in the elastic stage, increasing the pretension and anchorage length can effectively improve the working resistance. When the bolt is in the yield and strain-strengthening stages, increasing the pretension and anchorage length cannot effectively improve the working resistance. The influence of pretension and anchorage length on the ordinary and high-strength bolts is similar. The ordinary bolt's working resistance is approximately 25 kN less than that of the high-strength bolt. When pretension and anchorage length are considered separately, the best pretensions of the high-strength bolt in the middle of the roadway side and the roadway corner are 41.55 and 104.26 kN, respectively, and the best anchorage lengths are 1.54 and 2.12 m, respectively. The best anchorage length of the ordinary bolt is the same as that of the high-strength bolt, and the best pretension for the ordinary bolt in the middle of the roadway side and at the roadway corner is 33.51 and 85.12 kN, respectively. The research results can provide a theoretical basis for supporting the design of quadrilateral mining roadways.

为了解决刘庄煤矿150802胶带顺槽快速掘进过程中巷道帮角局部变形严重的问题,通过FLAC3D数值模拟软件确定了巷道异常区域的关键部位.结果表明:巷道在帮角处产生较大剪切应力集中区域,高帮角的剪切应力集中值大于低帮角,巷道顶板的位移变形量大于两帮,通过数值模拟确定巷道支护的第一关键部位是巷道顶板的两个肩窝,其次是顶板,最后是两帮.针对异常区域提出了优先支护的工艺顺序,并对巷道围岩采用"锚带网索"加强支护,锚杆(索)应力传感器现场实测应力是顶板两肩窝＞顶板＞两帮,与数值模拟结果相吻合,且巷道采用异常区域优先与"锚带网索"联合支护后,在巷道掘进110 m后,巷道围岩整体变形逐渐趋于稳定,巷道两帮的移近量和顶板的下沉量均低于45 mm,异常区域变形得到有效控制,实现了巷道平均日进尺16.8 m的快速掘进目标.

The mechanical model of the basic roof fracture structure is established on the basis of key block theory to study the roof breaking mechanism of gob-side entry retaining under roof cutting and pressure relief, and the analytical formula of roof support resistance is derived when the key block of the basic roof is stable. The influence of roof cutting angle and cutting height on roof support resistance is also analyzed. Determining the cutting seam parameters of the retained roadway roof is necessary to identify the support resistance of the roadway roof due to the correlation between the roof cutting parameters and the support resistance. Taking the II 632 haulage drift of the Hengyuan coal mine as the engineering background, FLAC3D numerical simulation is used in this paper to analyze the influence of different roof cutting angles and cutting heights on the surrounding rock structure evolution of retained roadways. Results show that the roof cutting angle and cutting height respond to the support resistance of the retained roadway roof, and the support resistance required by the roof increases with the roof cutting angle and cutting height. This condition ensures that the side roof of the gob can be cut off smoothly, and the support resistance required by the roof of retained roadways is within a reasonable range. Through theoretical and numerical simulation analysis, the reasonable roof cutting height of II 632 haulage drift is 8 m and the roof cutting angle is 15°. The theoretical analysis and numerical simulation results reveal that the required support resistance to maintain the stability of the roadway roof is 0.38 MPa. The supporting scheme of the roof of the II 632 haulage drift in the Hengyuan coal mine is then designed. Finally, the field industrial test is used for verification. The borehole imaging results show that the overall line of the retained roadway roof is small based on the description of field monitoring results. The deformation of the surrounding rock surface of the retained roadway is less than 100 mm, and the roadway is 40 m from the lagging working face. The deformation rate of surrounding rock decreases with the increase in distance from the working face. The integrity of the retained roadway roof is good, and the deformation of the surrounding rock is effectively controlled.

An overlying rock structure plays a key role in controlling the roof deformation of nonpillar gob-side entry retaining by roof cutting. On the bases of the actual geological conditions of II 632 Haulage Roadway at the Hengyuan coal mine, a similar three-dimensional simulation experiment of roof precutting is conducted. Thereafter, the caving characteristics and migration law of the roof strata in the strike and dip directions are obtained. Moreover, the roof of the retained roadway and key strata of the goaf can form a hinge structure of the key blocks. By monitoring the deformation of the surrounding rock and stress distribution of the roof, the skew deformation characteristics of roadway roof are obtained. By observing the borehole peeping technology, the roof subsidence near the goaf is determined to be greater than that of the solid coal side, and the roof subsidence of the gob-side entry retained by roof cutting is greater than that of the floor heave and two sides approaching. Results of the three-dimensional similar simulation experiment indicate that the mechanical structure model of the key block of the retained roadway roof is constructed, and the mechanical analytical solution of the required support resistance of the retained roadway roof is obtained. This study proposes the constant resistance and large deformation anchor cable reinforcement support method to control the roof deformation of the retaining roadway. Through engineering application, the maximum value of the roof and floor movement of the retained roadway is stable at approximately 650 mm. The retained roadway can meet the demand of the next mining face.