In situ observation of spalling process of intact rock mass at large cavern excavation

Accepted Manuscript In situ observation of spalling process of intact rock mass at large cavern excavation

Guofeng Liu, Xia-Ting Feng, Quan Jiang, Zhibin Yao, Shaojun Li PII: DOI: Reference:

S0013-7952(17)30029-7 doi: 10.1016/j.enggeo.2017.05.012 ENGEO 4575

To appear in:

Engineering Geology

Received date: Revised date: Accepted date:

7 January 2017 15 May 2017 25 May 2017

Please cite this article as: Guofeng Liu, Xia-Ting Feng, Quan Jiang, Zhibin Yao, Shaojun Li , In situ observation of spalling process of intact rock mass at large cavern excavation. The address for the corresponding author was captured as affiliation for all authors. Please check if appropriate. Engeo(2017), doi: 10.1016/j.enggeo.2017.05.012

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ACCEPTED MANUSCRIPT

In situ observation of spalling process of intact rock mass at large cavern excavation Guofeng Liua, Xia-Ting Feng a,1 , Quan Jiang a, Zhibin Yaob, Shaojun Lia a

State Key Laboratory of Geomechanics and Geotechnical Engineering, Institute of Rock and

Key Laboratory of Ministry of Education on Safe Mining of Deep Metal Mines, Northeastern

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b

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Soil Mechanics, Chinese Academy of Sciences, Wuhan 430071, China

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University, Shenyang 110819, China

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Corres ponding author. E-mail: [email protected]; [email protected]; Tel.:+86 13808671436;

ACCEPTED MANUSCRIPT Abstract This paper presents the results of in situ observation of rock spalling process of a large underground powerhouse, with 34 m in span and 88.7 m in height, during excavation layer by layer over two years. A method for in situ observation of rock spalling process has been established in pre-installed boreholes by using digital borehole camera. Observation results clearly show the whole process of rock fracturing and the associated spalling failure. Not only

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could the development of rock fracturing and spalling affected by the step-by-step excavations

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along the cavern's axis be captured, but also could the effect of the subsequent excavation layer by layer on the previous rock fracturing and spalling area be observed. The failure mechanism of

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rock mass surrounding the large cavern, which involves complex excavation sequences, could be better understood through the observation results. The difference of rock spalling behavior

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between large cavern and small tunnel is discussed. Influence of geostress, geological structures, rock mechanical properties, and layered excavation process on rock spalling process has been

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analyzed. The results in this study is instructive for rethinking the engineering design including the excavation and support at similar large underground caverns, as well as for the dynamic

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adjustment of excavation schemes and support optimization during the construction of a large cavern subjected to high geostress.

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Keywords

Large cavern excavation; Stress-induced fracturing; Rock spalling process; In situ observation;

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Digital borehole camera.

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Introduction

Rock spalling is a typical stress-induced failure near excavations in deep underground engineering, which often appears as rock splitting and flaky exfoliation. In the 1960s, Fairhurst and Cook (1966) had described this type of rock mass behavior and pointed out it is associated with the development of extension fractures parallel to compressive loading. Over the past few

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decades, spalling failure has been of great concern and attracted many scholars' extensive attention, especially in a series of deep underground experimental tunnels in the world, such as

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the Mine-by experiment tunnel in AECL’s Underground Research Laboratory, the Äspö pillar

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stability experiment (APSE) conducted in Hard Rock Laboratory in Sweden and the in situ experiment called Posiva’s Olkiluoto Spalling Experiment (POSE) conducted in Finland. The

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above typical cases have been widely studied in the field of rock mechanics and engineering, and many valuable findings have been obtained (Simmons, 1992; Martin and Simmons, 1993;

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Cundall et al., 1996; Martin, 1997; Read, 2004; Kaiser et al., 2000; Andersson, 2003; Cai et al., 2004; Staub et al., 2004; Martino and Chandler, 2004; Diederichs, 2005; Martin and Christiansson, 2009; Andersson et al., 2009; Hoek and Martin, 2014; Cai and Kaiser, 2014). As a

several meters in diameter.

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whole, current studies on rock spalling mainly focus on the small scale tunnels, usually with

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With the further development and utilization of underground space at great depths, more and more large underground caverns in deep engineering are being built or planned in

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highly-stressed rock mass. For example, the surge shaft under construction at Baihetan hydropower station in China has a span of 48 m and height of 100 m. For such cavern which has a high sidewall and large span, it must be excavated layer by layer and further divided into several excavation parts in each layer. Affected by the complex excavation sequences, rock mass is subjected to repeated stress adjustment, and therefore may suffer from specific progressive failure which differs from that of the small tunnel. As for a small tunnel, rock fracturing and the associated spalling near excavations always develop with tunnel advance, and finally almost remain unchanged when the working face moves far away. For instance, the spalling failure that occurred in Mine-by tunnel roof developed inward progressively with the step-by-step advance of working face and finally formed a 0.5 m deep notch geometry (Martin, 1997). However, as for

ACCEPTED MANUSCRIPT a large cavern, the spalling could proceed over a lengthy period of time during not only the step-by-step advancing of working faces along the cavern axis but also the downward excavation layer by layer. Considering the completed underground powerhouse cavern with a span of 28.9 m and a height of 68.8 m at Jinping I hydropower station in China as an example (Li et al., 2009), rock mass near excavation was often subjected to the stress-fracturing and spalling failure during the first layer of excavation (i.e., the roof of the cavern), and the failure gradually

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stopped after the systematic rock support completed. When the excavations proceeded to the

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third layer, reaching an excavation height of up to 22 m, cracks in the shotcrete layer appeared in the downstream roof of the cavern, and developed progressively as excavation continued.

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Finally, the shotcrete layer in the downstream roof, almost along the whole cavern, suffered from cracking. As a consequence, the loading on a few rockbolts and anchor cables exceeded

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ultimate capacity, and the steel arch support was bended. In fact, these hazards were directly caused by the progressive rock fracturing as the cavern was excavated downwards. The results

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of ultrasonic testing in borehole showed that the depth of excavation damaged zone (EDZ) in the roof was within 2 m after the first layer of excavation, but increased to approximately 4 m

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after the third layer of excavation completed, even reached up to 6-7 m in some boreholes after the fifth layer of excavation. It seems that the failure around large cavern excavations in brittle

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rock mass under high geostress brings new requirements for engineering design. Since the design should match the rock’s behavior, it is very essential to know the behavior of brittle rock mass at such large cavern excavation, especially to know how the rock fracturing and the

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associated induced failure occurs and develops in the whole excavation process, i.e., to determine the relationship between rock fracture initiation, propagation and termination and the excavation process.

Digital borehole camera technology is a visualized observation method. Rock fractures through the borehole wall can be clearly visible with the naked eye by using a borehole camera, then the location, width and orientation of rock fracture can be obtained quickly and accurately. The technology has been widely applied in geotechnical engineering field (Lau et al., 1987; Miyakawa et al., 2000; Schepers et al., 2001; Wang et al., 2002; Lahti, 2004). It has also been used to observe the fracturing development at tunnel excavation, contributing to the establishment of the EDZ characteristics (Yuji, 1983; Li et al., 2012a, b). In 2013, the ISRM

ACCEPTED MANUSCRIPT suggested method for rock fractures observations using a borehole digital camera was proposed (Li et al., 2013). Although the observation using borehole camera technology is localized, it is more specifically suitable for observing rock fracturing process from surface to inner rock mass in EDZ. Study on observation of rock fracturing evolution and the associated spalling process under complex excavation schemes around a large cavern has not been carried out. This study attempts to reveal typical spalling process of intact rock mass at large cavern

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excavation, based on proposing an in situ observation method using digital borehole camera

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technology. The proposed in situ observation method has been conducted for over two years in a large underground powerhouse, with an excavation span of 34 m and an excavation height of

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88.7 m, at a hydropower station under construction in the southwest of China. A typical case has been analyzed, and observation results clearly show the progressive development of rock

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fracturing and spalling during the whole excavation process. According to this study, the stress-induced failure mechanism of brittle rock mass under complex excavation process could

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be better understood. Engineering design involving the excavation and rock support at large cavern excavation could thus be more conveniently realized. Dynamic adjustment of excavation

In situ observation of spalling process

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2

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schemes and support optimization could be more scientific-oriented.

2.1 Observation purpose

In general, micro tensile fractures in the rock will appear, extend, and intersect typically in

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parallel to the maximum stress direction under no or small confinement (Tapponier and Brace, 1976; Horii and Nemat, 1986). As a result, the multiple tensile fractures can lead to macro splitting face. During the cavern excavation, the tangential stress parallel to excavation boundary rapidly increases while the normal stress decreases dramatically, even to zero, which could make surrounding rock suffer from the stress-fracturing approximately parallel to excavation boundary. As the cavern working face advances, rock mass will be split into instable slices or slabs by intersected fractures (see Fig.1), and be developed into spallling failure. Spalling is induced by the propagation and coalescence of stress-fracturing from surface to internal rock. The key for better understanding rock spalling failure at large cavern excavation is to record the development characteristics of rock fracturing with the step-by-step excavation

ACCEPTED MANUSCRIPT along cavern axis and subsequent layer-by-layer excavation downwards. That is, the high-risk area at underground excavation should be regularly observed. Based on the comparison of fractures observed at different times, the development of stress-fracturing including initiation, propagation, and termination that occurred in surrounding rock, and the associated spalling failure process could be obtained.

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2.2 Observation apparatus There are two main types of digital borehole camera (Li et al., 2013), one of which is a digital

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optical televiewer, and the other of which is a digital panoramic borehole camera. The imaging

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principles of the two types of borehole camera are the same, and the full 360 degree image of the borehole wall can be obtained by using them. However, the way of image capturing and

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transmission for the two types is different. As for a digital optical televiewer, an annulus image is captured and transmitted every time, and then the annulus images are spliced together into a

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whole piece of image along with time, while a whole piece of image could be captured and transmitted at one time by using a digital panoramic borehole camera. In this context, a digital

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panoramic borehole camera is used for in situ observation of rock fracturing and associated spalling process. It mainly consists of probe, depth measuring device, leading sheave, borehole

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optical imager, cables, tripod, and alternative measuring rods for horizontal or inclined boreholes. The highest circumferential accuracy of which is 0.1-0.2 mm. Details on the principles, components and operations of the borehole camera system are obtainable from the

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relevant reference (Li et al., 2013).

2.3 Layout schemes of observation boreholes In order to ensure the stability during construction and operation of a large cavern, as well as functional requirements, some small auxiliary tunnels are always designed and excavated earlier around the large underground cavern, such as the anchorage-monitoring tunnels, drainage tunnels, etc., which are very common in underground hydropower engineering. Making the most of these pre-excavated tunnels distributed near the concerned cavern will be very helpful to the layout of observation boreholes. As illustrated in Fig.2, the flow chart of the dynamic layout for observation boreholes focusing on spalling failure process around large underground

ACCEPTED MANUSCRIPT excavations is thus demonstrated. As to the positional distribution of observation boreholes, the principle is that boreholes should always be distributed in the high-risk spalling area. For a large carven, the occurrence of new spalling is possible with the downward excavation layer by layer. Therefore, a risk assessment work should be carried out before the excavation of each layer to predict the high-risk spalling area, considering the engineering geology information, geostress field, designed excavation

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schemes and the analysis of spalling failure cases in earlier excavation or nearby pre-excavated

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caverns. Then suitable positional distribution of the observation boreholes could be determined. It is noted that there is a common conclusion that stress-induced spalling always occurs at the

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excavation profile parallel to or at small angle with the orientation of maximum far-field principal stress in the cross section of the tunnel, which has been proved in many deep tunnels

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under high geostress (Stephansson et al., 1989; Hoek et al., 1995; Haimson and Lee, 1995, Read 2004; Martin and Christiansson 2009; Jiang et al., 2013). This conclusion is also applicable in

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large caverns. Besides, in the case that some unexpected new spalling risks emerge during excavation, new observation boreholes could also be added. In a word, the observation

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boreholes should be arranged dynamically with the excavation of the cavern. There are two ways to arrange boreholes, i.e., pre-drilled boreholes and immediate-drilled

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boreholes after excavation of the tunnel or cavern. Pre-drilled borehole refers to the borehole prepared before the excavation and is close to the concerned risk area in cavern. It includes two cases. One is to drill boreholes through the concerned risk area by means of the pre-excavated

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auxiliary tunnels around the cavern, as illustrated in Fig.3a, and the other is to use the early excavated space of the cavern to drill boreholes through the concerned risk area (in Fig.3b and Fig.3c). The advantages of pre-drilled observation boreholes are obvious, namely the whole process of fracturing in the rock mass around the borehole before and after the excavation could be observed. In addition, the type of boreholes shown in Fig.3a could also be used as long-term observation boreholes for assisting in observing the stability of such a large carven during its construction and operation period. Immediate-drilled borehole means that the observation borehole is directly drilled in the concerned area of the cavern immediately after the adjacent excavation, as illustrated in Fig.3d. Only the fracturing development of rock mass after excavation could be observed, however, such type of observation borehole is relatively

ACCEPTED MANUSCRIPT flexible to be arranged. Once new spalling risk arises during cavern excavations, observation boreholes could be added timely. The ways of arranging boreholes need to be determined according to the specific engineering characteristics, and any one or all of above ways could be adopted. 2.4 Frequency of observation

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Generally, rock fracturing and spalling are usually concentrated in the stage of stress redistribution near excavations. Thus, when excavation is approaching the observation borehole,

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the frequency of observation needs to be increased. As a diagrammatic example shown in

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Fig.3b and Fig. 3c, the observation could begin when the working face advances to a position, from where there is still a certain distance (x) to the observation section, and may be stopped

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when the working face passes through the observation section for a certain distance (y), until the rock fracturing near the observation section temporarily stops growing. The values of x and

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y mainly depend on the range of the excavated affected zone along the cavern axis, which may change for different engineering. It should be noted that the excavation process of S2 is used

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just as an example to illustrate the frequency of observation. And the frequency of observation is the same with that in the excavation process of S2 during the subsequent excavation process

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Case study

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3.1 Introduction

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of S3, S4, etc. in Fig.3b.

The in situ observation method proposed above has been applied in a large cavern at a hydropower station located in the southwest of China, which has a dimension of 453 m ×34 m × 88.7 m (length × span × height). The cavern, still under construction, has a burial depth of 260-330 m and a full length of 453 m (from chainage K0-72 to chainage K0+381), with the axis direction of N20°E. Additionally, two same tunnels, with an arch cross section of 4.5×5 m, were excavated at about 27m above the powerhouse in parallel, which are used as anchorage-monitoring tunnels. On the one hand, high-strength anchor cables could be made by the two tunnels to reinforce the roof of the large powerhouse. On the other hand, some equipment could be expediently buried through the two tunnels for monitoring or observing

ACCEPTED MANUSCRIPT the rock mass behavior of the powerhouse. A three-dimensional representation of the underground powerhouse engineering conditions is shown in Fig.4. The rock mass in the powerhouse cavern is mainly composed of basalt of the Upper Permian Emeishan Formation, relatively intact and belonging to class II- or class III 1 according to the Standard for Engineering Classification of Rock Masses, the National Department of Technical Monitorial Affairs and the Ministry of Construction, PRC, GB50218-94 (Feng and Hudson 2011) .

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In this standard, rock mass is classified according to uniaxial compressive strength Rc, integrity Kv,

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plus the influence of underground water, initial stress state, and orientation of main weak structural planes. The type of basalt has characteristics of both hardness and brittleness (σc ≈

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120-150 MPa, σt ≈ 10-18 MPa). Main geological structures involved in this large cavern are also shown in Fig.4, therein, F717 , F720 and F721 are three steep-inclined faults, respectively, and WIZ

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No. 2 and WIZ No.3152 are two weak interlayer zones, respectively. The geostress is mainly dominated by tectonic stress, and the horizontal stress is greater than the vertical stress. The

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first and second principal stresses are almost horizontal and the third is vertical. Maximum principal stress is approximately within 20-30 MPa, with the orientation of N30°~50°W and the

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dip angle of 6°-15°. Intermediate principal stress is approximately within 13-20 MPa and the minimum principal stress is nearly the gravity stress, approximately 8-12 MPa.

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The powerhouse is being excavated downward layer by layer from top to bottom. Because the height of the cavern changes along with the cavern axis (as shown in Fig. 4 ), the numbers of excavation layers along the axis are correspondingly different, and the design height of each

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layer sometimes can be adjusted by back analysis, however, ranges from 4 meters to 10 meters usually. By August 2016, the excavation of the top four layers ( Ⅰ-Ⅳ) of the cavern had been completed, and the detailed excavation sequences of which are represented in Fig.7a. Due to the large excavation span, the stability of rock mass in the roof of the cavern is of great concern. With respect to layer Ⅰ, a central pilot tunnel (Ⅰ1 ) was first excavated (12 × 10 m). After the systematic support in the roof of the central pilot tunnel was completed, the floor of the central pilot tunnel was excavated downwards by 1 m ( Ⅰ2 ). Then, central pilot tunnel was expanded by 6 m on both upstream and downstream sides ( Ⅰ3 ). Next, the floor on both sides

ACCEPTED MANUSCRIPT was again excavated downwards by 2.6 m (Ⅰ4 ), when the systematic support in part Ⅰ3 was accomplished. Finally, the excavation of the first layer ( Ⅰ) was finished by excavating 5 m towards both sides. Drilling and blasting method was adopted, and the length of each excavation round along the cavern axis was generally between 2 m and 5 m. It is noted that, excavation could proceed on both sides at the same time, in addition, there were several working faces

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advancing along the cavern axis on each side. The rock support design in the cavern roof is shown in Fig.5.

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Affected by the stress redistribution, the rock mass in the roof of the cavern was generally

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subjected to stress-induced failure during excavation. In particular, the cavern from chainage K0+310 to chainage K0+350 suffered from serious fracturing and brittle spalling failures, as

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shown in Fig.6a. Dominated by the orientation of maximum stress on the cross section of the cavern, spalling failure mainly occurred on the upstream side of the cavern roof, with the failure

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depth of 10-50cm generally, even up to 100 cm, which had repeatedly threatened the construction safety and caused the failure of the rock support system. Fig. 6b shows a typical

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rock spalling during excavation of layer Ⅰ.

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3.2 Layout schemes of in situ observation boreholes Considering that rock spalling frequently occurred in the roof from chainage K0+310 to chainage K0+350 during excavation of the central pilot tunnel ( Ⅰ1 ), the roof of cavern from chainage

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K0+310 to chainage K0+350 would be likely to suffer from spalling failure during subsequent excavation. In order to continuously observe the spalling process of rock mass in this area during subsequent excavations, the observation schemes were put forward after the excavation of the central pilot tunnel. The key points are as follows: Steep-inclined boreholes were drilled downward from anchorage-monitoring tunnels to the roof-top of the cavern before the excavation of part Ⅰ3 , with 110 mm in diameter and approximately 26.5 m in depth, as shown in Fig.7a. There were totally four observation boreholes downwards which were symmetrically distributed on both sides (upstream side and downstream side) of the roof-top, on section K0+320 and K0+330, respectively (in Fig.7b). It is

ACCEPTED MANUSCRIPT noted that the bottom of such boreholes remained a distance of approximately 1.2 m from the excavation boundary. That is because that this type of boreholes could also be used for ultrasonic testing, which need be operated with water injection. In addition, four slightly inclined boreholes, with 110 mm in diameter and 22 m in length, were symmetrically distributed in the corner of the roof on both sides, on section K0+320 and K0+330,

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respectively. Such boreholes were drilled after central pilot tunnel excavated, and passed through the unexcavated zone (Ⅰ3 and Ⅰ5 ) to reach the final surrounding rock (in Fig. 7a).

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These symmetrical boreholes could be used to contrastively analyze the rock fracturing on both

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upstream and downstream roof of the cavern, therein, the boreholes distributed in the upstream roof-top and the downstream corner of cavern roof should be paid high attention,

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since they are arranged in high-risk spalling area, according to the orientation of σmax on the cross section of the cavern.

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3.3 Case analysis

By August 2016, the in situ observation work ha d last for over two years. A typical case is

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presented as below to illustrate the observation process and the corresponding results. On June 25, 2014, the working face Ⅰ3 on the upstream side (Ⅰ3 -U for short) advanced from north to south

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and reached section K0+335, and the pre-drilled observation borehole K0+330-0-U located in the upstream roof was 5m ahead the working face, as illustrated in Fig.8. During subsequent

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excavation, the behavior of rock mass around borehole K0+330-0-U was highly concerned and observed constantly, which was not only dominated by the step-by-step advancing of working faces along the cavern axis in layer Ⅰ, but also affected by the subsequent excavations downwards layer by layer. The detailed process of rock spalling behavior is illustrated as follows: Rock behavior associated with the advancing of working faces in layer Ⅰ Fig.9a-9k show the borehole images of rock fracturing within 4 m originated from excavation boundary in borehole K0+330-0-U observed at different times during the excavation of layer Ⅰ of the cavern. Fig.10 shows the corresponding photos of rock spalling failure process around borehole K0+330-0-U. Furthermore, combined with the excavation and rock support

ACCEPTED MANUSCRIPT information near the observation borehole, Fig.11 illustrates the development process of observed stress-fracturing and the associated spalling failure in excavation damaged zone around borehole K0+330-0-U, which is analyzed in detail as follows: On June 25, 2014, i.e., before the working face Ⅰ3 -U advanced to chainage K0+330, borehole image indicated that rock mass in the borehole K0+330-0-U was almost intact, as shown in

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Fig.9a. It can be seen that there were few macroscopic fractures except two slight spalling veins on the borehole wall.

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On June 29, 2014, working face Ⅰ3 -U passed through observation section K0+330 and advanced

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to chainage K0+329, as shown in Fig.10a, the excavation boundary around borehole K0+330-0-U was relatively smooth (corresponding to No.0 in Fig.11). Rock mass near the bottom of borehole,

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i.e., 1.2-1.4 m from the excavation boundary, could not be observed (in Fig.9b) due to the blocking caused by the fractured rocks hanging on the borehole wall. It can be inferred that

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stress-fracturing of surrounding rock may occur within the depth of 1.4 m after the adjacent excavation.

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On July 2, 2014, the working face Ⅰ3 -U advanced to chainage K0+325, 5 m away from section K0+330. There were new fractures occurring through the borehole (in Fig. 9c) in comparison

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with the previous observation result, and the maximum depth of fracturing was 1.75 m. Before the steel-fiber shotcrete layer with 5 cm in thickness was applied in the surrounding rock from chainage K0+330 to chainage K0+325, the rock mass near the borehole K0+330-0-U, from

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chainage K0+330 to chainage K0+328, was subjected to spalling failure, with 30 cm in depth, as shown in Fig.10b (corresponding to No.1 in Fig.11). On July 3, 2014, the working face Ⅰ3 -U advanced to K0+323, 7 m away from section K0+330. New fractures appeared, and the maximum depth of rock fracturing in borehole sharply increased to 2.40 m (in Fig.9d). On July 9, 2014, working face Ⅰ3 -U advanced to K0+319, 11 m away from section K0+330. The rock fracturing in borehole continued to develop and increase to 2.56 m in depth (in Fig.9e). The borehole from 1.20m to 1.80m in depth could not be observed due to new blocking near the bottom.

ACCEPTED MANUSCRIPT On July 12, 2014, the working face Ⅰ3 -U stopped advancing, remaining 13 m away from observation section K0+330. Then the pre-stressed rockbolts support (Φ32 mm, L=9 m, T=100KN, @1.2 m×1.2 m) in the area around borehole K0+330-0-U started to be applied and completed on July 16. Afterwards, the depth of rock fracturing around borehole developed slowly (in Fig.11). On July 16, 2014, the maximum depth of rock fracturing in borehole increased

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to 2.73 m (in Fig.9f). On August 3, 2014, the working face Ⅰ3 -U advanced to K0+308, 22 m away from section K0+330.

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Borehole K0+330-0-U was re-drilled and extended to the excavation boundary. Borehole image

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showed that the fractures in the borehole had been very developed, especially in the rock mass near the excavation boundary (in Fig.9g), and small rockfalls of the borehole wall occurred due

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to the coalescence of some fractures. The depth of fracturing slightly increased to 2.78 m. In addition, the width of existing rock fractures located at different distance from the excavation

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boundary, also increased in comparison with previous observation results (in Fig.11). On August 7, 2014, the rock mass from chainage K0+330 to chainage K0+337 suffered from new spalling

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failure, as shown in Fig.10c (corresponding to No.2 in Fig.11), with the failure depth of approximately 0.6 m, and parts of rock support were therefore destroyed.

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On August 16, 2014, reinforcing fabric (Φ8 mm, @15 cm×15 cm.) was applied to the concerned area around borehole K0+330-0-U. By this time, the working face Ⅰ3 -U had advanced to K0+290, 40 m away from section K0+330, the influence of which on observation section K0+330 became

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little. While the working face on the downstream side (Ⅰ3 -D) was excavated to section K0+320, and was advancing closer and closer towards section K0+330 in later, which caused the damaged extent of existing rock fractures around borehole K0+330-0-U to increase again. As shown in Fig.11, although the maximum depth of rock fracturing stopped rising, the width of existing rock fractures still increased, resulting in further development in spalling failure. On August 25, the depth of rock spalling around borehole K0+330 -0-U increased to 0.8m, as shown in Fig.10d (corresponds to No.3 in Fig.11). Compared to before, the surface of surrounding rock was more fractured and became rougher. On September 2, 2014, the working face Ⅰ3 -D advanced to section K0+360. The maximum depth

ACCEPTED MANUSCRIPT of rock fracturing in borehole K0+330-0-U was still 2.78 m (in Fig.9h), while the width of existing rock fractures still slightly increased (in Fig.11). On September 16, 2014, the working face on the downstream side (Ⅰ4 -D) began to advance on both directions along cavern axis from section K0+325, and another working face on downside side (Ⅰ5 -D) also followed to advance on both directions from section K0+325. On September 18,

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the working face (Ⅰ4 -D and Ⅰ5 -D) passed through section K0+330 and advanced to section K0+335.

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Affected by the frequent excavations nearby, although the maximum depth of observed fracturing in borehole K0+330-0-U was still 2.78 m (in Fig.9i), the damaged extent of existing

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rock fracturing around the borehole was intensified further, and the depth of spalling failure reached up to 1 m, as shown in Fig.10e (corresponding to No.4 in Fig.11). The rock mass around

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borehole K0-330-0-U had been generally fractured. As the working face (Ⅰ4 -D and Ⅰ5 -D)

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continued to advance step by step, the extent of rock fracturing in the concerned area became more and more serious. Consequently, the steel mesh reinforcement was obviously squeezed by the fractured rocks.

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On October 3, 2014, the depth of rock fracturing still remained 2.78 m (in Fig.9j), however, the rock mass around borehole K0+330-0-U were generally fractured, and the rock support seemed

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increasingly unsustainable. To ensure the construction safety and long-term stability of the cavern, the fractured rock mass was cleaned away and the destroyed rock support in this area

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was removed, as shown in Fig.10f (corresponding to No.5 in Fig.11). It was found that the depth of unstable loosened rock mass caused by fracturing had reached up to 1.8 m, forming an asymmetric v-shaped notch. Rock support with high intensity was remade after the unstable rock mass were cleaned away (Pre-stressed bolting: Φ32, L= 9 m, T=100KN, @1.2 m×1.2 m. Shotcrete sprayer layer: 20 cm thick. Steel mesh: Φ8 mm, @15 cm×15 cm). Fig.10g shows the v-shaped notch geometry around borehole K0+330-0-U after the renewed support completed. Afterwards, during the advancing of working faces on upstream side (Ⅰ4 -U and Ⅰ5 -U), on the one hand, the excavation length in one round was decreased compared to before, on the other hand, rock support reinforcement was requested to be applied quickly after excavation, which was conductive to the reduction of rock fracturing and spalling failure. By the end of November,

ACCEPTED MANUSCRIPT 2014, the excavation of layer Ⅰ had been nearly completed. Then anchor cables were applied in the cavern roof (approximately 25.5-29.5 m in length), and secondary steel-fiber shotcrete layer was carried out later (approximately 15 cm in thickness). The latest borehole image indicated that the rock fractures in borehole K0+330-0-U almost remained unchanged (in Fig.9k). Obviously, during the cavern excavation of layer Ⅰ, the surrounding rock in the upstream roof

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near section K0+330 shows a typical progressive failure due to repeated influences of the excavations of different working faces nearby. For better understanding the above relationships

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between spalling development and the excavation progress, the results of in-situ investigation

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and observation obtained at different times are clearly presented by corresponding sketches, as shown in Fig.12, which could visually reflect the progress of several working faces, the update of

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rock support conditions, the propagation of rock fracturing close to borehole, and the development of spalling failure profile around borehole K0+330-0-U.

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Rock behavior associated with the downward excavations layer by layer The cavern from chainage K0+310 to chainage K0+350 was divided into four layers in total, by

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August, 2016, the excavation of the cavern from chainage K0+310 to chainage K0+350 had been completed. Fig.9l-9n show the observed images of rock fracturing in borehole K0+330-0-U after

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each layer of the cavern excavated, and the development process of rock fracturing and the associated failure in the downward excavation layer by layer was illustrated in Fig.11, which was

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described as follows:

By June 2015, the excavation of layer II, with a height of 4.1 m, had been completed. The rock fracturing observed after the excavation of layer II (in Fig.9l) was almost unchanged compared to the results observed after layer I (in Fig.9k) except that the bottom of the borehole was blocked and out of observation. While the displacement of upstream cavern roof at chainage K0+328 started to increase, as shown in Fig.11. (The displacement data corresponding to the monitored point M-1.5 in the multipoint extensometer MzcK0+328-3 is provided in Fig.11, and the point M-1.5 means the monitored point is 1.5 meters from the surface of cavern wall. Note that this extensometer (in Fig.8) was installed after the excavation of pilot tunnel of the cavern, but it was broken in the subsequent excavation of layer I and was restored until the excavation

ACCEPTED MANUSCRIPT of layer I was completed. Therefore, the displacement presented in Fig.11 was not the total displacement of upstream cavern roof at chainage K0+328, but the displacement increments after the excavation of layer I) By January 2016, the excavation of layer Ⅲ, with a height of 11 m, had been completed. As shown in Fig.9m, there were several new smallish rock fractures occurring through the borehole

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wall, and the depth of rock fracturing extended to 3.5 m, in addition, the range of blocking near the bottom of the borehole became larger. Meanwhile, the displacement of Point M-1.5

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increased rapidly during the excavation of layer III, and the displacement increment reached to

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approximately 5 mm, as shown in Fig.11. Meanwhile, it was found that some slight cracks in steel-fiber shotcrete layer around borehole K0+330-0-U were arising from chainage K0+325 to

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chainage K0+335 during the excavation of layer Ⅲ, as shown in Fig.13. It was inferred that the fracturing extent of rock mass in the EDZ around borehole K0+330-0-U developed further. To

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ensure construction safety, the reinforcing fabric was then applied again on the upstream side of the cavern roof when the excavation of layer Ⅲ was nearly completed.

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By August 2016, the excavation of layer Ⅳ, with a height of 5.9 m, had been completed. Observation results (in Fig.9n) indicated that there was no increase in the depth of rock

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fracturing, while the width of existing rock fractures increased, as described in Fig.11. In addition, the displacement of point M-1.5 increased again by the end of the excavation of layer

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IV. There were also new rockfalls occurring on the borehole wall due to the coalescence of some fractures. Meanwhile, in situ investigation indicated that the range and extent of cracking in the steel-fiber shotcrete layer around borehole K0+330-0-U were further intensified from chainage K0+310 to chainage K0+340, some of which had developed into shotcrete flaking, as shown in Fig.14. 3.4 Preliminary analysis of rock spalling mechanism 3.4.1 Influence of geological conditions Fig.15 shows the basic geological information of the cavern from chainage K0+270 to chainage K0+380, including the main geological structures, dominant sets of joints, litholog ical distribution, rock fracturing and spalling area during excavation, etc.

ACCEPTED MANUSCRIPT From aspect of the regional geologic structures, spalling failure mainly located on the overlapped area between the footwall of the steep-inclined fault F720 and the upper wall of the slightly-inclined weak interlayer zone No.2. The rock mass exposed along with fault or ISN are always of poor quality and cataclastic due to the structural movement in historical period, usually damaged by the forms of structure-induced failures, such as collapse or rockfall, but not spalling failure. However, the fault or WIZ could affect the geostress state regionally. Previous

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researchers had proved that the magnitude of the geostress within a certain zone near the fault

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may vary obviously, based on analyzing the geostress data of underground engineering involving the URL in Canada, Forsmark nuclear power plant, eight hydropower stations and two mines

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from India (Martin and Chandler, 1993b; Stephansson and Angman, 1986; Sengupta, 1999). The variation of geostress near the geological structure is complicated, but usually could be divided

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into two aspects: the first is that the geostress presents discontinuity from the upper wall to the footwall of the fault, the other is that the geostress within a certain zone near the fault may

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increase. According to the distribution of rock failures in cavern, it is inferred that the geostress in the region from chainage K0+310 to chainage K0+350 seems at a relatively high level affected

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by the fault F720 and the WIZ No.2. In addition, among the boreholes in the cavern, only the borehole K0+330-0-U suffered from core discing phenomenon, as shown in Fig.16.

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Rock mass from chainage K0+310 to chainage K0+350 is almost intact and of high quality, belonging to II-class surround rock according to the rock mass rating system of BQ. In general, stress-induced failures tend to easily occur in such rock mass.

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There are four types of rocks in cavern, including cryptocrystalline basalt, porphyritic basalt, amygdaloidal basalt and breccia lava, respectively. Rock spalling mainly occurred in the porphyritic basalt. The laboratory test results indicate that the porphyritic basalt is hard and brittle and the average uniaxial compressive strength (UCS) is approximately 135MPa which is the lowest among the four rock types. 3.4.2 Influence of geostress The above Fig.6 and 8 show the relationship between geostress state and spalling area on the cross section of the cavern. The far-field maximum and minimum principal stress (σ1 and σ3 ), in the plane of analysis, are approximately 25-30 MPa and 13-15 MPa, respectively. As the part I 3

ACCEPTED MANUSCRIPT on the upstream side at chainage K0+330 was excavated, the maximum tangential elastic stress in the upstream roof would reach approximately 50-55 MPa. As a result, the stress level index SL =σθ / UCS goes up to approximately 0.37-0.42, in the range of 0.3-0.5 which is widely considered as the in situ rock spalling strength, supported by Martin et al. (1999), Andersson et al. (2009), and many others. As the excavation continued, the maximum tangential elastic stress in the

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upstream roof would slightly increase, eventually goes up to approximately 60-65 MPa after the excavation of layer IV. It is noted that the above analysis result was obtained without taking the

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influence of the irregularities of the excavation boundary, which was regarded as an “apparent”

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result actually (Cai and Kaiser, 2014). 3.4.3 Influence of rock mechanical property

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The laboratory experimental results indicated that basalt rock is characterized by the density of 3

approximately 2.8 g/cm , the UCS of 120-150 MPa, the tensile strength of 10-15 MPa, the

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longitudinal wave velocity of 5,100-5,600 m/s, the elasticity modulus of 35-48 GPa, and the poisson ratio of 0.2-0.3. Fig.17 shows a stress-strain curve of porphyritic basalt sample,

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representing typical characteristics of tensile failure under compressive loading. Similarly, rock mass near excavations, with high tangential stress and very low confined pressure, is sensitive

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to such tensile fracturing and failure. Rock spalling pieces near chainage K0+330 was selected for analyzing the fracture mechanism by using scanning electron microscope. The scanning plane is the new fracturing plane which is approximately parallel to the excavation boundary (in

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Fig.18a). Scanning results (in Fig.18b) show that the crystal surfaces are smooth and distributed like the conchoidal shape, and the edge of fracture is sharp, showing transgranular fracture. There are a few scattered rock debris at the lower-right corner of the scanned plane which has some abrasive veins, however, no parallel scratches. As a result, it can be inferred that the scanned surface is characterized by tensile fracture. Overall, the high geostress, hard brittle rocks and rock mass structure constitute the potential geologic environment of spalling formation, and excavation activities cause the freedom of the spalling failure eventually. The orientation of geostress is the reason why the severity of rock spalling on different parts of the cavern is different. Especially, the continuous step-by-step and layer-by-layer excavation process is the essential which result s in the progressive failure, in such

ACCEPTED MANUSCRIPT a cavern with large span and high sidewall. 4

Discussion As for a small tunnel, rock fracturing and the associated spalling often occurs after

excavation unloading, then develops with the advance of working face and almost remains unchanged finally when the working face move s far away, such as the progressive spalling

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occurred in Mine-by test tunnel (Martin, 1997). However, the progressive development of rock fracturing and the associated spalling failure in such a large cavern, which involved complex

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excavation sequences, are different from those of the small tunnel, as described by a

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diagrammatic drawing in Fig.19. As for a large cavern, rock spalling to excavations would proceed over a lengthy period of time and may develop into a large scale failure finally. On the

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one hand, the rock mass in the immediate zone surrounding an excavation where the confinement is low would be subjected to stress-fracturing and possible brittle failure

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(Corresponding to phase 1 in Fig.19). On the other hand, affected by the advance of other nearby working faces, surrounding rock of early excavated zone would suffer from stress adjustment again, which could intensify the fracturing and associated rock failure

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(Corresponding to phase 2, phase 3, etc. in Fig.19). What’s more, as the cavern excavated downward layer by layer (Corresponding to phase n, n-1, etc. in Fig.19), the depth-span ratio of

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cavern shape changes from <1 to >1, which could further intensify the extent of stress concentration in specific area (Area A in Fig.19, or such as the upstream side of cavern roof in

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the above case). Consequently, as excavation continues, the depth or extent of rock fracturing and the associated failure in this area often increases by a multistep development process. Then, the geometric bulking caused by the inside rock fracturing and failure could impose large radial deformations on the support system and could result in the rock support losing effectiveness gradually, such as the cracking or breakout of steel-fiber shotcrete layer, rock-bolting failure, etc. as shown in Fig.20. During the excavations of this large cavern mentioned above, the rock behaviors to the excavations along the whole upstream side of cavern roof were characterized by such multistep development process. When this large cavern was excavated to layer IV, large number of shotcrete cracks appeared on the upstream cavern roof, and the accumulative length of these cracks, which were interruptedly distributed along the cavern axis, accounted for

ACCEPTED MANUSCRIPT approximately 70 percent of the total length of the cavern. According to Martin’s findings (1999), the depth of brittle failure in the tunnel can be approximated by a linear relationship given as 𝐷𝑓 𝑎

= 1.25

σ𝑚𝑎𝑥 σ𝑐

− 0.51(±0.1) (1)

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where σmax is the maximum tangential stress in the surrounding rock; σ c is the uniaxial compressive strength; Df is the failure depth; a is the tunnel radius or effective tunnel radius.

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It is noted that this linear relationship is obtained by analyzing those case histories

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associated with small tunnels. Assuming that there is a small tunnel which has the same magnitude of σ𝑚𝑎𝑥 and σ𝑐 with the large cavern presented above, i.e., σmax = 65 MPa

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(Corresponding to the maximum stress around borehole K0+330 -0-U after the cavern excavation of layer IV) and σc =135 MPa, respectively. Using Eq.1, the range of rock failure depth

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(𝐷𝑓 ) in this assumed tunnel could be estimated, i.e., 𝐷𝑓 = (0-0.192) a. According to the observation results obtained from borehole camera, as shown in Fig. 3.3, the fractures within 2.8 depth were dense and continuously distributed along the borehole wall after the excavation

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of layer IV completed. Actually, the rock mass within 2.8 m had been broken by those connected fractures, and was just restrained by rock support, however, the shotcrete layer was

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consequently squeezed and cracked. Therefore, the failure depth of rock mass could be determined as 2.8 m (D = 2.8 m). By the way shown in Fig.19, the effective radius (a) of the

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cavern involved in four layers is 20.8 m. The ratio of the failure depth to effective cavern radius at chainage K0+330 is 0.135 accordingly, in the range from 0 to 0.192 which is obtained by Eq. (1). Noted that the depth of rock failure associated with this large cavern was measured under the condition of being supported by high-strength measures. Anyhow, the above analysis is just a discussion according to the observation results from this cavern. It still needs to be verified by further studies of more case histories from those large caverns. When such a large cavern excavated, the fracturing may extend gradually along with the step-by-step and layer-by-layer excavations, and eventually stabilizes at a certain degree. The possible ultimate depth or extent of rock fracturing should be estimated at the early stage of the cavern excavation, by means of numerical simulation based on back analysis. Rock support,

ACCEPTED MANUSCRIPT therefore, can be more scientific-oriented and determined earlier, and matches the rock mass behavior as the excavation of the cavern goes on. 5

Conclusions

In order to study the specific characteristics of spalling process of intact rock mass in a large cavern, which involved complex excavation sequences, an in situ observation approach to reveal

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rock spalling process using digital borehole camera technique is determined. Several observation layout schemes that can be flexibly arranged are provided, by which the rock

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fracturing and the associated progressive development of spalling failure with the complex

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excavations of a large cavern can be clearly observed and traced.

The in situ observation practice has been successfully applied over two years in a large

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underground powerhouse, with an excavation span of 34 m and an excavation height of 88.7 m, at a hydropower station from southwest of China. Observation results indicated that rock

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fracturing and the associated spalling failure appeared quickly after excavation, and progressively developed with the advance of the current working face. Meanwhile, it was

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intensified gradually by the advance of other nearby working faces in the same layer of the cavern, and tended to become stable temporarily when these working faces gradually moved

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away. Afterwards, the subse quent downward excavations layer by layer would lead to the increase of the depth or extent of the early stress-fracturing and the associated spalling failure. Rock spalling to excavations of such a large cavern can proceed over a lengthy period of time

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and may develop into a large scale failure finally. The results could contribute to study the specific characteristics of rock mass behavior at large cavern excavation under high geostress, and could also provide some new ideas for the engineering design including the excavation and rock support at similar large underground caverns. Acknowledgements The authors gratefully thank the financial support from the National Natural Science Foundation of China under Grant no. 41320104005 and 11232014. In particular, the authors would also express their sincere thanks to prof. Q.X. Fan, Prof. Y.L. Fan and Mr X.P. Duan for their kind help in the in-situ investigation and technical support from China Three Gorges Project Corporation.

ACCEPTED MANUSCRIPT The authors would also like to thank Dr. S.Q. Duan and Dr. S.F. Pei who gave support and assistance during the observation of boreholes in the hydropower station project. References Andersson, J.C., 2003. Äspö Pillar Stability Experiment, Feasibility study. SKB Report IPR-03-01. Andersson, J.C., Martin, C.D., Stille, H, 2009. The Äspö pillar stability experiment: part II—rock

Journal of Rock Mechanics and Mining Sciences 46(5), 879-895.

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mass response to coupled excavation-induced and thermal-induced stresses. International

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Cai, M., Kaiser, P.K., Tasaka, Y., Maejima, T., Morioka, H., Minami, M., 2004. Generalized crack

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initiation and crack damage stress thresholds of brittle rock masses near underground excavations. International Journal of Rock Mechanics and Mining Sciences 41(5), 833–847.

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Cai, M., Kaiser, P.K., 2014. In-situ rock spalling strength near excavation boundaries. Rock Mechanics and Rock Engineering 47(2), 659-675.

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Cundall, P.A., Potyondi, D.O., Lee, C.A., 1996. Micromechanics-based models for fracture and breakout around the mine–by tunnel. In: Martino, B., Martin, C.D. (Eds.), Proceedings of the Canadian Nuclear Society International Conference on Deep Geological Disposal of

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Nuclear Waste, Winnipeg. Canadian Nuclear Society, Toronto, pp. 113–122. Diederichs, M.S., 2007. The 2003 canadian geotechnical colloquium: mechanistic interpretation

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and practical application of damage and spalling prediction criteria for deep tunnelling. Canadian Geotechnical Journal 44(9), 1082-1116.

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Fairhurst, C., Cook, N.G.W., 1966. The phenomenon of rock splitting parallel to the direction of maximum compression in the neighborhood of a surface. Proceedings of the first congress on the international society of rock mechanics, Lisbon, pp. 687-692. Feng Xia-Ting, Hudson John 2011 Rock Engineering Design, CRC Press. Haimson, B.C., Lee, C.F., 1995. Estimating in situ stress conditions from borehole breakouts and core disking—experiment results in granite. In: Sugawara, K., Matsuki, K. (Eds.), Proceedings of the International Workshop on Rock Stress Measurement at Great Depth, Eighth ISRM Congress, The Workshop, pp. 19–24. Hoek, E., Kaiser, P.K., Bawden, W.F., 1995. Support of underground excavations in hard rock. A.A. Balkema, p. 215.

ACCEPTED MANUSCRIPT Hoek, E., Martin, C.D., 2014. Fracture initiation and propagation in intact rock-a review. Journal of Rock Mechanics and Geotechnical Engineering 6(4), 287-300. Horii, H., Nemat, N.S., 1986. Brittle failure in compression: splitting, faulting and brittle ductile transition. Philosophical Transactions of the Royal Society of London, Series A 319, 337–374.

spalling veins: a case study. Engineering Geology 152(1), 38-47.

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Jiang, Q., Feng, X.T., Chen, J., Huang, K., Jiang, Y.L., 2013. Estimating in-situ, rock stress from

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Kaiser, P.K., Diederichs, M.S., Martin, C.D., Sharp, J., Steiner, W., 2000. Underground works in hard rock tunnelling and mining. In:Keynote lecture at GeoEng2000, Technomic Publishing

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Co.,Melbourne, Australia, pp. 841–926.

Lau, J.S.O., Auger, L.F., Bisson, J.G., 1987. Subsurface fracture surveys using a borehole television

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camera and acoustic televiewer. Canadian Geotechnical Journal 24, 499–508. Lahti, M., 2004. Digital borehole imaging of the boreholes KR24upper part and PH1 at Olkiluoto,

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Posiva Oy. Working Report 2004-28, pp. 21.

Li, Z.K., Zhou, Z., Tang, X.F., Liao, C.G., Hou, D.Q., Xing, X.L., Zhang, Z.Z., Liu, Z.G., Chen, Q.H.,

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2009. Stability analysis and considerations of underground powerhouse caverns group of Jinping Ⅰ hydropower station. Chinese Journal of Rock Mechanics and Engineering

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28(11), 2167-2175 (in Chinese).

Li, S.J., Feng, X.T., Li, Z.H., Chen, B.R., Zhang, C.Q., Zhou, H., 2012a. In situ monitoring of rockburst nucleation and evolution in the deeply buried tunnels of Jinping II hydropower

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station. Engineering Geology 137–138, 85–96. Li, S.J., Feng, X.T., Li, Z.H., Zhang, C.Q., Chen, B.R., 2012b. Evolution of fractures in the excavation damaged zone of a deeply buried tunnel during TBM construction. International Journal of Rock Mechanics and Mining Sciences 55(10), 125–138. Li, S.J., Feng, X.T., Wang, C.Y., Hudson, J.A., 2013. Isrm suggested method for rock fractures observations using a borehole digital optical televiewer. Rock Mech and Rock Engineering 46(3), 635-644. Martin, C.D., Chandler, N.A., 1993b. Stress heterogeneity and geological structures. International Journal of Rock Mechanics & Mining Sciences & Geomechanics Abstracts 30(7), 993-999.

ACCEPTED MANUSCRIPT Martin, C.D., 1997. Seventeenth canadian geotechnical colloquium: The effect of cohesion loss and stress path on brittle rock strength. Canadian Geotechnical Journal 34(5), 698-725. Martin, C.D., Kaiser, P.K., Mccreath, D.R., 1999. Hoek-brown parameters for predicting the depth of brittle failure around tunnels. Canadian Geotechnical Journal 36(1), 136-151. Martin, C.D., Christiansson, R., 2009. Estimating the potential for spalling around a deep nuclear waste repository in crystalline rock. International Journal of Rock Mechanics and Mining

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Sciences 46(2), 219-228.

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Martino, J.B., Chandler, N.A., 2004. Excavation-induced damage studies at the underground research laboratory. International Journal of Rock Mechanics and Mining Sciences

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41(8),1413-1426.

Miyakawa, K., Tanaka, K., Hirata, Y., Kanauchi, M., 2000. Detection of hydraulic pathways in

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fractured rock masses and estimation of conductivity by a newly developed TV equipped flowmeter. Engineering Geology 56(1–2), 19–27.

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Read, R.S., 2004. 20 years of excavation response studies at AECL's underground research laboratory. International Journal of Rock Mechanics and Mining Sciences 41(8), 1251-1275.

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Schepers, R., Rafat, G., Gelbke, C., Lehmann, B., 2001. Application of borehole logging, core imaging and tomography to geotechnical exploration. International Journal of Rock

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Mechanics and Mining Sciences 38(6), 867–876. Sengupta, S., 1999. Influence of geological structures on in-situ stresses. Journal of Orthopaedic Research 17(2), 192-199.

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Simmons, G.R., 1992. The Underground Research Laboratory Room 209Excavation Response Test—A summary report. Atomic Energy of Canada Limited Report AECL-10564, COG-92-56.

Staub, I., Andersson, J.C., Magnor, B., 2004. Äspö pillar stability experiment, geology and mechanical properties of the rock in TASQ. SKB report R-04-01, Stockholm Stephansson, O., Angman, P., 1986. Hydraulic fracturing stress measurements at Forsmark and Sidsvig, Sweden. Bulletin of the Geological Society of Finland, no 58, Part I, pp. 307-333. Stephansson, O., Savilahti, T., Bjarnason, B., 1989. Rock mechanics of the deep borehole at Gravberg, Sweden. In: Fourmaintraux, D., Maury, V. (Eds.), International Symposium Rock at Great Depth, vol. 2. A.A. Balkema, pp. 863–870.

ACCEPTED MANUSCRIPT Tapponier, P., Brace, W.F., 1976. Development of stress induced microcracks in Westerly granite. International Journal of Rock Mechanics and Mining Science and Geomechanics Abstracts 13, 103–112. Wang, C.Y., Law, K.T., Sheng, Q., Ge, X.R., 2002. Borehole camera technology and its application in the Three Gorges project. In: Proceedings of the 55th Canadian geotechnical and 3rd joint IAH-CNC and CGS groundwater specialty conferences, Niagara Falls, Ontario, pp.

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601–608.

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Yuji, K., 1983. Observation of crack development around an underground rock chamber by

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borehole television system. Rock Mechanics and Rock Engineering 16(2), 133–142.

ACCEPTED MANUSCRIPT Caption of Figures Fig.1 Diagrammatic drawing of the stress state and corresponding stress-fracturing in rock mass near underground excavations Fig.2 The basic flow chart of the dynamic layout of boreholes for observing spalling process around large cavern excavations

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Fig.3 The sketch of in situ layout schemes of observation boreholes for rock spalling process at large cavern excavation：(a) Pre-drilled observation boreholes arranged by the pre -excavated

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auxiliary tunnels around the cavern; (b) Pre-drilled observation boreholes arranged by the pre-excavated space of the cavern; (c) Planar diagram of A-A section in Fig.3b for illustration on

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the frequency of in situ observation of rock spalling during cavern excavation; (d) The observation boreholes arranged immediately after adjacent excavation.

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Fig.4 The spatial location of the underground powerhouse engineering and involved geological structures.

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Fig.5 The cross-section profile of the rock support design in the cavern roof Fig.6 The Stress-fracturing and spalling failure phenomenon during cavern excavation of layer I

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(a) The shape of the cavern roof affected by rock fracturing and spalling from chainage K0+310 to chainage K0+350, (b) A typical spalling failure on the upstream side of the cavern roof.

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Fig.7 In situ layout schemes of observation boreholes in the roof of the cavern from chainage K0+310 to chainage K0+350 (a) Layout of observation boreholes on the cross-section (Note: K0+XXX-0-2 represents the No. of the observation borehole, therein, K0+XXX is the chainage of

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observation section, -0 represents the roof-top, -1 represents the corner of roof, -U represents upstream side and -D represents downstream side), (b) The planar distribution of the observation boreholes.

Fig.8 The potential spalling area and corresponding observation borehole K0+330-0-U at the cavern excavation on June 25, 2014. Fig.9 Observation images of rock fracturing within 4 m originated from excavation boundary in borehole K0+330-0-U measured at different times during cavern excavation from layer I to layer IV (Note: "June. 25, 2014" is the date of observation. "Ⅰ3 -U" represents that the nearest working face to observation section K0+330 is the "Ⅰ3 " on the upstream side of the cavern; "U"

ACCEPTED MANUSCRIPT represents upstream side; "D" represents downstream side; "-5m" represents that the nearest working face is 5 m away and is moving towards the observation section K0+330; " -" represents the related working face has not yet passed through observation section K0+330; and "+" represents the nearest working face has passed though and moved gradually away from observation section K0+330).

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Fig. 10 Photos of rock spalling failure process around the observation borehole K0+330-0-U during the excavation of layer Ⅰ: (a) Working face Ⅰ3 -U just passed through section K0+330,

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(b) Spalling failure occurred in the area from chainage K0+331 to chainage K0+329, with a depth

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of 30 cm, (c)Rock mass from chainage K0+337 to chainage K0+329 suffered from spalling failure, with a depth of 60 cm, (d) The depth of spalling failure around borehole K0+330-0-U increased

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to 80 cm, (d) The depth of spalling failure increased to 100 cm, (e) The depth of spalling failure increased to 175 cm, (f) The final v-shaped notch geometry around borehole K0+330-0-U after

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completion of the renewed support.

Fig. 11 Development of the depth and width of rock fracturing and the associated spalling failure occurred around the borehole during cavern excavation from layer I to layer IV. (Note:

labels of photos in Fig.10).

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the numbers of a-n correspond to the labels in Fig.9, and the numbers of 0-5 correspond to the

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Fig. 12 Progressive development of rock spalling failure around the observation borehole K0+330-0-U under the varying excavation scenes. (The numbers of profile 0-5 correspond to the

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labels in Fig. 10 and 11. Additionally, only the rock fractures which were observed from borehole K0+330-0-U and excavation surface were presumably depicted in above figures). Fig.13 Cracking of shotcrete layer around observation borehole K0+330-0-U during the cavern excavation of layer Ⅲ

Fig.14 More serious cracking of shotcrete layer around borehole K0+330 -0-U during the cavern excavation of layer Ⅳ Fig.15 The planar distribution of geological information of the cavern wall from chainage K0+270 to chainage K0+380. Fig.16 The core discing phenomenon in basalt cores from the borehole K0+330-0-U. Fig.17 Typical experiment result of porphyritic basalt under uniaxial compression.

ACCEPTED MANUSCRIPT Fig.18 Scanning electron microscopy (SEM) results of the rock pieces caused by spalling failure around borehole K0+330-0-U (a) In situ sampling location and scanning plane (b) Typical SEM image of failure plane associated with rock spalling (Magnification: 800 on left and 2000 on right). Fig.19 Comparison of the spalling development process to excavations between the small tunnel and the large cavern. (Note: the influence of humidity and temperature on spalling

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development is not considered).

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Fig.20 Failure of those earlier-completed rock support caused by the stress-fracturing and spalling of rock mass that occurred during subsequent excavation of the large cavern (a)

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Cracking of shotcrete layer; (b) Squeezing of rock support.

ACCEPTED MANUSCRIPT σ1

σ3

σ2

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Engi neering geology i nformation Geostress field Des igned excavation s chemes Ana l ysis of the failure cases i n earlier exca vation or nearby pre-excavated ca verns

SC

   

PT

Fig.1 Di a grammatic drawing of the stress s tate a nd corresponding stress-fracturing i n rock ma s s nea r underground exca va ti ons

NU

Predi cti ng the pote nti a l s pa l l i ng ri s k a rea before the exca va ti on of l a yer i

Dri l ling observation boreholes from pre-exca va ted tunnel s

MA

nea rby or the ea rl y exca va ted s pa ce of the ca vern

ED

Keeping observation of s pall i ng ri s k a rea from borehol es wi th the exca va ti on proces s

EP T

i=i+1

Duri ng the ca vern exca va ti on of l a yer i

AC C

Exca va ti ng next l a yer

Occurrence of new s pa l l i ng ri s k？ No

Yes

Addi ng new obs erva ti on boreholes (drilling boreholes di rectl y i n s pa l l i ng a rea or from pre-excavated tunnel s ) (Alternative)

Ca vern excavation compl eted？ Yes End

Fig.2 The ba sic flow chart of the dynamic l ayout of borehol es for obs ervi ng s pa l l i ng proces s a round l a rge ca vern exca va ti ons

ACCEPTED MANUSCRIPT

Pre-excavated tunnel

Observation borehole Potential spalling risk area Pre-excavated tunnel Observation borehole

Potential spalling risk area

Unexcavated

PT

Pre-excavated tunnel

Designed boundary between two excavation layers

SC

Underground cavern

（a ） Observation borehole Potential spalling risk area S3

Observation borehole

S4 S5

Potential spalling risk area

EP T

Early working face Advance (S2 )

ED

Underground cavern (during the excavation of part S2 )

(b) (c)

y

S1

Current working face

x

A

NU

S2

A-A

Surrounding rock

MA

A

RI

Observation borehole

Potential spalling risk area

AC C

Observation boreholes

Excavated

Underground cavern (two excavation parts has been completed)

Potential spalling risk area

(d) Fig.3 The sketch of in situ la yout schemes of observa tion boreholes for rock spalling process a t large ca vern exca va tion ：(a ) Pre-drilled observa tion boreholes a rranged by the pre-exca va ted auxiliary tunnels a round the ca vern ; (b) Pre-drilled observa ti on boreholes a rranged by the pre -exca va ted spa ce of the ca vern; (c) Plana r diagra m of A-A section in Fi g.3b for illustra tion on the frequency of in situ observa tion of rock spalling during ca vern exca va tion; (d) The observa tion boreholes a rranged i mmediatel y after adja cent exca va tion.

ACCEPTED MANUSCRIPT

F721 F717 Anchorage-monitoring tunnels WIZ No.3152

PT

Underground powerhouse

SC

RI

F720

WIZ No.2

Fig.4 The s pa ti a l l oca ti on of the underground powerhous e engi neeri ng a nd i nvol ved geol ogi ca l s tructures .

Prestressed anchor cable, T = 2000 kN, L= 25 m, longitudinal space = 3.6 m

80.5°

53° 44° Central Pilot tunnel

40°

AC C

Prestressed rockbolt, T = 100 kN L = 9 m, longitudinal space = 1.2m, rectangular staggered

ED

EP T

25°

Grouted rockbolt, L = 6 m Prestressed rockbolt, T = 100 kN, L = 9 m, @1.2×1.2m, rectangular staggered

75.6°

75.6°

80.5°

44°

MA

Prestressed anchor cable, T =2000 kN, L=25.5–29.5m, longitudinal space = 3.6 m

53°

2# anchorage-monitoring tunnel

NU

1# anchorage-monitoring tunnel

R = 21 m

25° 40° Grouted rockbolt, L= 9 m @1.2×1.5 m

100°

Steel-fiber shotcrete layer, thinkness = 5 cm Steel-fiber reinforced shotcrete with mesh, thickness = 15cm

Fig.5 The cros s -s ecti on profi l e of the rock s upport des i gn i n the ca vern roof

ACCEPTED MANUSCRIPT

Downstream side

Upstream side

Spalling area

Ⅰ1

Ⅰ3

Ⅰ3

PT

Ⅰ2

(a) (b)

RI

Fig.6 Stres s-fracturing and spalling failure phenomenon during ca vern excavation of layer I (a) The shape of the ca vern roof a ffected by rock fra cturi ng a nd s palling from chainage K0+310 to cha inage K0+350, (b) A typi cal s palling failure on the upstream side of the

SC

ca vern roof.

Anchorage-monitoring tunnels

Downstream side

2#

K0+XXX-0-U

K0+XXX-0-D

12m

Ⅰ5 Ⅱ2 Ⅲ6 Ⅲ2 Ⅲ4 Ⅲ5

Ⅰ4

Ⅰ1 Ⅰ2 Ⅱ1 Ⅲ1 Ⅲ3

Ⅳ1 4m

Ⅳ1

Ⅰ5

613.6m 611.0m 606.9m

Ⅱ2 Ⅲ2 Ⅲ6 601.4m Ⅲ4 Ⅲ5 595.9m 590.0m

EP T

Upstream side

Ⅰ3 Ⅰ4

624.6m

K0+XXX-1-D

ED

Ⅰ3

6m 5m

23m

4m

F20

Downstream side

614.6m

Porphyritic basalt

Ⅰ3 Ⅰ5

MA

290° 5m 6m 22m

Ⅰ5 Ⅰ3 Ⅰ1

26.5m

K0+XXX-1-U

NU

1#

K0+300 K0+310 K0+320 K0+330 K0+340 Upstream side

624.6m

614.6m K0+350

Observation borehole in the roof-top of cavern Observation borehole in the corner of roof Multipoint extensometer in the cavern roof Spalling area during the excavation of central pilot tunnel F Fault Dominant sets of joints Central axis in the roof of the cavern 720

(a) (b)

Fig. 7 In situ l a yout s chemes of observation boreholes i n the roof of the cavern from chainage K0+310 to cha inage K0+350 (a ) Layout

AC C

of obs ervation boreholes on the cross-section (Note: K0+XXX-0-2 represents the No. of the observation borehole, therein, K0+XXX is the cha inage of observation section, -0 represents the roof-top, -1 represents the corner of roof, -U represents upstream side and -D represents downstream side), (b) The planar distribution of the observa tion boreholes.

PT

ACCEPTED MANUSCRIPT

RI

Mzc0+328-3 K0+330-0-U M-11

Potential spalling area

SC

M-6.5 M-3.5 M-1.5

290°

I5

NU

I1

I3

II2

Upstream side

I2

I4 II1

I3 I4

I5 II2

Downstream side

AC C

EP T

ED

MA

Fig. 8 The potential spalling a rea and corresponding observation borehole K0+330-0-U a t the cavern exca va ti on on June 25, 2014.

ACCEPTED MANUSCRIPT a) June. 25, 2014; I 3-U, -5m; before excava tion (The borehole bottom remains a dis tance of 1.2 m from the designed excava tion boundary)

1.2m Spalling of borehole wall

4.0

3.8

3.6

3.4 3.2 3.0 2.8 2.6 2.4 2.2 2.0 1.8 1.6 Distance originated from the excavation boundary of the cavern

1.4

Designed excavation boundary

1.2

PT

b) June.29, 2014; I 3-U, +1m; a fter excavation

1.2m

3.8

3.6

3.4 3.2 3.0 2.8 2.6 2.4 2.2 2.0 1.8 1.6 Distance originated from the excavation boundary of the cavern

1.4

SC

4.0

RI

Blocking near the bottom of borehole, hard to observe

NU

c) Jul y.2, 2014; I 3-U, +5m

1.2

Blocking, hard to observe

4.0

3.8

3.6

MA

New fracture

3.4 3.2 3.0 2.8 2.6 2.4 2.2 2.0 1.8 1.6 Distance originated from the excavation boundary of the cavern

1.2

ED

d) Jul y.3, 2014; I 3-U, +7m

1.4

4.0

3.8

3.6

EP T

New fracture

3.4 3.2 3.0 2.8 2.6 2.4 2.2 2.0 1.8 1.6 Distance originated from the excavation boundary of the cavern

1.4

1.2

AC C

e) Jul y.9, 2014; I 3-U, +11m

Blocking, hard to observe

Previous fracture

Blocking, hard to observe

New fracture

4.0

3.8

3.6

3.4 3.2 3.0 2.8 2.6 2.4 2.2 2.0 1.8 1.6 Distance originated from the excavation boundary of the cavern

1.4

1.2

f) Jul y.16, 2014; Ⅰ3-U, +13m

New fracture 4.0

3.8

3.6

3.4 3.2 3.0 2.8 2.6 2.4 2.2 2.0 1.8 1.6 Distance originated from the excavation boundary of the cavern

Blocking, hard to observe 1.4

1.2

Actual excavation boundary

ACCEPTED MANUSCRIPT g) Augus t.3, 2014; I 3-U, +22m (Re-drilling the borehole, borehole was extended to the excavation boundary) A deflected bolt-hole by mistake Small rockfall New fracture 4.0

3.8

3.6

3.4

3.2

3.0

2.8 2.6 2.4 2.2 2.0 1.8 1.6 1.4 1.2 1.2 Distance originated from the excavation boundary of the cavern

1.0 1.0

0.8 0.8

0.6 0.6

0.4 0.4

0.2 0.2

00

3.8

3.6

3.4

3.2

3.0

2.8 2.6 2.4 2.2 2.0 1.8 1.6 1.4 1.2 Distance originated from the excavation boundary of the cavern

1.0

Spalling of surrounding rock

0.8

NU

i ) September.18, 2014; I 4-D, +5m; I 5-D, +5m

3.8

3.6

3.4

3.2

3.0

2.8 2.6 2.4 2.2 2.0 1.8 1.6 1.4 1.2 Distance originated from the excavation boundary of the cavern

ED

j) October.3, 2014; I 4-D, +25m; I 5-D, +20m

3.8

3.6

3.4

3.2

3.0

1.0

2.8 2.6 2.4 2.2 2.0 1.8 1.6 1.4 1.2 Distance originated from the excavation boundary of the cavern

1.0

AC C

k) November.27, 2014; I 5-U, +28m, the excavation of layer I was nearly completed

4.0

3.8 3.6 3.4 3.2 3.0 2.8 2.6 2.4 2.2 2.0 Distance originated from the excavation boundary of the cavern

1.75

l ) Jul y, 2015, After the excavation of layer Ⅱ

Blocking, hard to observe 4.0

0.2

0

3.8 3.6 3.4 3.2 3.0 2.8 2.6 2.4 2.2 2.0 Distance originated from the excavation boundary of the cavern

0.8

0.6

0.4

0.2

0

0.4

0.2

0

Spalling of surrounding rock

EP T

4.0

0.4

Spalling of surrounding rock

MA

4.0

0.6

SC

4.0

RI

PT

h) September.2, 2014; I 3-D, +30m

1.75

0.8

0.6

ACCEPTED MANUSCRIPT m) Ja nuary, 2016, After the excavation of layer ⅡI Due to the oil spot on the probe

New fracture 4.0

Blocking, hard to observe

3.8 3.6 3.4 3.2 3.0 2.8 2.6 2.4 2.2 2.0 Distance originated from the excavation boundary of the cavern

1.75

PT

n) August, 2016, After the excavation of layer I V

3.8 3.6 3.4 3.2 3.0 2.8 2.6 2.4 2.2 2.0 Distance originated from the excavation boundary of the cavern

1.75

SC

4.0

RI

Blocking, hard to observe

Fig.9 Observa tion images of rock fracturing wi thin 4 m ori ginated from exca va tion bounda ry i n borehole K0+330-0-U measured at

di fferent ti mes during ca vern exca vation from la yer I to layer IV (Note: "June. 25, 2014" is the da te of observa tion. " Ⅰ3-U" represents

NU

tha t the nea res t working fa ce to observa tion section K0+330 is the " Ⅰ3" on the upstream side of the ca vern; "U" represents upstream side; "D" represents downstrea m side; "-5m" represents that the nea rest working fa ce is 5 m awa y and is moving towa rds the observati on section K0+330; "-" represents the related working fa ce has not yet passed through observa tion secti on K0+330; and

AC C

EP T

ED

MA

"+" represents the nearest working face has passed though and moved gradually a way from observation s ection K0+330).

ACCEPTED MANUSCRIPT (a ) June.29, 2014

PT

K0+330-0-U

0

RI

(b) Jul y.2, 2014

MA

NU

SC

K0+330-0-U

1

ED

(c) Augus t. 7, 2014 (d) August. 25, 2014

EP T

K0+330-0-U

AC C

K0+330-0-U

2

(e) September. 18, 2014

K0+330-0-U

4

3

ACCEPTED MANUSCRIPT (f) October. 3, 2014 (g) October. 8, 2014

Supported againregion Re-supported

PT

K0+330-0-U

5

Fig.10 Photos of rock s palling failure process a round the observation borehole K0+330-0-U during the excavation of layer Ⅰ: (a )

RI

Worki ng face Ⅰ3 -U just passed through section K0+330, (b) Spalling failure occurred i n the area from chainage K0+331 to cha inage K0+329, wi th a depth of 30 cm, (c)Rock mass from chainage K0+337 to cha inage K0+329 s uffered from spalling failure, with a depth

SC

of 60 cm, (d) The depth of spalling failure a round borehole K0+330-0-U increased to 80 cm, (d) The depth of s palling failure i ncreased to 100 cm, (e) The depth of s palling failure increased to 175 cm, (f) The final v-s haped notch geometry a round borehole

AC C

EP T

ED

MA

NU

K0+330-0-U a fter completion of the renewed s upport.

ACCEPTED MANUSCRIPT Layer I

Layer II

-30m I3-D +30m -5m I5-D +30m +30m

I3-U

-5m

Shotcrete layer Rockbolts

-10m I4-D +30m

Layer III

Layer IV

0m I5-U +30m -10m I4-U +30m Cracking of shotcrete layer

Reinforcing fabric Remaking rock support

Displacement of rock mass in upstream cavern roof at chainage K0+328 (Point M-1.5, Mzc0+328-3)

20

0

2014/8/24

2 2014/7/15

2014/8/4

Date (Year/Month/Day)

20 20

50 50

2.732.73 m from boundary m fromexcavation excavation boundary m fromexcavation excavation boundary 2.532.53 m from boundary 0.75 m from excavation boundary 0.75 m from excavation boundary

99

2014/11/12

2015/7/14

2014/12/2

2016/1/20

66 33 2016/8/25

0

2.27 m from excavation boundary 2.27 m from excavation boundary 1.70 m from excavation boundary 1.70 m from excavation boundary

NU

40 40

l

100

30 30

40

2014/10/23

50 2014/8/24 2014/9/13 2014/10/3 2014/10/23 2014/11/12 2014/12/2 0

10 10

30

2014/10/3

k

150

4

3

250 200

5

b

a 0

j

2014/9/13

c

2014/6/251

0

2014/8/4

i

n

Displacement(mm)

2014/7/15 de

h

m

PT

100 100 5050

g

f

12 12

RI

150 150

(mm) Fracturing width (mm) Fracturingwidth

50

MA

Fig.11 Devel opment of the depth a nd width of rock fracturing and the associated spalling failure occurred a round the borehole duri ng cavern excavation from layer I to layer IV. (Note: the numbers of a -n correspond to the labels in Fig.9, a nd the numbers of 0-5

EP T

ED

corres pond to the l abels of photos in Fig.10).

AC C

Fracturing width (mm)

250 250 2014/6/25 200 200 0

10

Rock support information in the area around borehole K0+330-0-U 400 Maximum depth of rock fracturing observed from borehole K0+330-0-U 350 Depth of rock spalling in the area around borehole K0+330-0-U Occurrence of shotcrete cracking around borehole K0+330-0-U 300

400 400 350 350 300 300

SC

Depth (cm) Depth

(cm)

I3-U The distance variation from working face (I -U) to section K0+330 3

ACCEPTED MANUSCRIPT Spalling risk area Borehole New fracture

Borehole

K0+330-0-U

0

Unexcavated

C

290° E

B I1 I5

A

E

I1 I3 I5

I5

I2

I4

I4

F

A

Advance

I1

D E F

B

K0+330-0-U

C D

E

Unexcavated

I5

K0+360

K0+350

K0+340

K0+330

K0+320

K0+310

K0+300

K0+290

New fracture

I3

F

I5

K0+370

K0+360

I1 I3 I5

I2

B C

I5 I3 I1

D E F

Advance Excavated

I5

F

B C

K0+310

The plane graph of excavation schedule

(c) Jul y.2, 2014 (d) August.7, 2014

I5

K0+300

K0+290

F

Unexcavated

I3 I1

I5 K0+320

I3 I5 I4

F

A

K0+330-0-U

K0+330

290°

Observation section (K0+330)

Concrete spraying

K0+340

K0+290

0

I2 I4

A

E K0+350

K0+300

E

I5

D

K0+360

D

I3

I3

K0+370

K0+310

2

I1

Observation section

Unexcavated

K0+320

B

Observation section (K0+330) A

K0+330

C

I3

AC C

A

K0+340

1

290°

I4

I4

Concrete spraying

Borehole

E

B

K0+350

K0+330-0-U

0

EP T

D

Excavated

ED

Previous fracture

C

Observation section

Advance

MA

(a) June.25, 2014 (b) June.29, 2014

1

F

The plane graph of excavation schedule

The plane graph of excavation schedule

Borehole

Unexcavated

I3 I1

Excavated

I3

K0+370

I5

SC

C

I3

NU

B

Observation section

Unexcavated

I5

I4

Observation section (K0+330) A

I5

I3

I2

I4

Observation section (K0+330) A

290°

B

I3

I3

D

PT

D

RI

C

Observation section K0+330-0-U

Excavated

Advance

Concrete spraying, Pre-stressed rockbolt support

I3

Advance

I5

K0+370

K0+360

K0+350

K0+340

K0+330

K0+320

K0+310

The plane graph of excavation schedule

K0+300

K0+290

ACCEPTED MANUSCRIPT

4

3

3

Borehole

2

1

C

290° E

B

I1 I5 A

E

I1 I3

I5

I2

I5

I4

I4

F

A

I1

D E

F

B

Advance

K0+330-0-U

C

Excavated Concrete spraying, Pre-stressed rockbolt support,Reinforcing fabric Advance

E

I5

K0+370

K0+360

K0+350

K0+340

K0+330

K0+320

K0+310

K0+300

K0+290

The plane graph of excavation schedule

4 3

I3 I5

I2

Unexcavated

K0+320

K0+310

K0+300

K0+290

I5

F

Observation section

K0+330-0-U

Making rock support again

I1

I4 Excavated

Advance

I3 I5

Advance K0+360

K0+330

I3

AC C

A

F

K0+340

290°

Observation section (K0+330)

K0+370

K0+350

0

I4

I4

A

E

K0+360

Advance

E

I1

D

I5

K0+370

I4 Excavated

ED

D

B

I3

Advance

I3

EP T

C

C

K0+330-0-U

Concrete spraying, Pre-stressed rockbolt support,Reinforcing fabric

2

1

I5

F

Observation section

MA

5

B

F

The plane graph of excavation schedule

(e) Augus t.25, 2014 (f) September.18, 2014

Borehole

I3 I1

D

I3

Unexcavated

I5

SC

I3

NU

C

A Observation section

Unexcavated

I5

I4

Observation section (K0+330)

A I5

I3

I2

I4

Observation section (K0+330)

B

290°

B

I3

I3

D

PT

D

2 0

1

RI

C

Borehole

0

K0+350

K0+340

K0+330

K0+320

K0+310

The plane graph of excavation schedule

K0+300

K0+290

Profile

Date (year/month/day)

Failure depth on section K0+330 (m)

0

2014/6/29

0

1

2014/7/2

0.30

2

2014/8/7

0.40

3

2014/8/25

0.45

4

2014/9/18

0.60

5

2014/10/3

1.75

(g) October.3, 2014 Fig.12 Progres sive development of rock spalling failure around the observation borehole K0+330-0-U under the va rying excavation s cenes. (The numbers of profile 0-5 correspond to the labels in Fig. 10 a nd 11. Additionally, only the rock fractures which were obs erved from borehole K0+330-0-U a nd excavation surface were presumably depicted i n a bove fi gures).

ACCEPTED MANUSCRIPT

a

1m Crack

a

b

K0+330-0-U

b

PT

1m

RI

Cracks

Fig.13 Cra cki ng of s hotcrete l a yer a round obs erva ti on borehol e K0+330-0-U duri ng the ca vern exca va ti on of l a yer Ⅲ

a

b

NU

SC

1m

MA

1m

c

Cracks

b

K0+330-0-U

c

EP T

ED

1m

AC C

d

a

d

1m

Fig.14 More s eri ous cra cki ng of s hotcrete l a yer a round borehol e K0+330-0-U duri ng the ca vern exca va ti on of l a yer Ⅳ

ACCEPTED MANUSCRIPT 590.0m

Downstream side

596.0m 606.9m

cryptocrystalline basalt

Breccia lava

WIZ No.2

Ⅳ

ⅠⅠⅠ ⅠⅠ

611.0m

Breccia lava Ⅰ

F720 624.6m

Porphyritic basalt

Porphyritic basalt

Ⅰ 611.0m

PT

Breccia lava

606.9m

cryptocrystalline basalt Breccia lava WIZ No.2

590.0m Upstream side

RI

596.0m

ⅠⅠ

ⅠⅠⅠ Ⅳ

Central axis on top of the cavern

Dominant sets of joints WIZWeak Interlayer

zone

NU

Rock fracturing and spalling region

SC

K0+270 K0+280 K0+290 K0+300 K0+310 K0+320 K0+330 K0+340 K0+350 K0+360 K0+370 K0+380

Lithologic boundary

F720

Fault

AC C

EP T

ED

MA

Fig.15 The pl a na r di s tri buti on of geol ogi ca l i nforma ti on of the ca vern wa l l from cha i na ge K0+270 to cha i na ge K0+380.

ACCEPTED MANUSCRIPT

PT

Core discing

RI

Fig.16 The core di s ci ng phenomenon i n ba s a l t cores from the borehol e K0+330-0-U.

140

σ1/MPa

100

εε1 1 εε3 3 εεvv

40

NU

-0.010

60

20

-0.005

0 0.000 ε

MA

-0.015

80

SC

120

0.005

0.010

ED

Fig.17 Typi ca l experi ment res ul t of porphyri ti c ba s a l t under uni a xi a l compres s i on .

EP T

Scanning plane

(a) (b)

AC C

Fig.18 Scanning electron mi cros copy (SEM) results of the rock pieces caused by spalling failure a round borehole K0+330-0-U (a ) In situ sampling loca tion and s canning plane (b) Typi cal SEM image of failure plane associa ted wi th rock spalling ( Magnifi cation: 800 on l eft a nd 2000 on ri ght).

ACCEPTED MANUSCRIPT

Phase 1: Advance of the excavation face adjacent to the concerned area Phase 2-n: Advance of subsequent excavation faces nearby the concerned area

Rf a

Phase n

A large cavern Phase 3

Area A

Phase 2 Phase 1

A small tunnel

SC

Excavation Progress

Area B

PT

Area B Df a

RI

Depth or extent of rock failure

Area A Df

Fig.19 Compa rison of the spalling development process to exca vati ons between the small tunnel and the large cavern. (Note: the

EP T

ED

MA

NU

i nfluence of humidity a nd temperature on spalling development is not considered).

(a) (b)

Fig.20 Failure of those ea rlier-completed rock support caused by the stress-fra cturing and spalling of rock mass tha t occurred during

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s ubsequent excavation of the large ca vern (a) Cra cking of shotcrete layer; (b) Squeezing of rock s upport.

ACCEPTED MANUSCRIPT Highlights 

An in situ observation method for rock spalling process at large cavern excavation is determined; Rock spalling process during the whole excavation process of large cavern is obtained;

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Difference of rock spalling behavior between large cavern and small tunnel is realized

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