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Comprehensive evaluation method of stope stability and its application in deep metal
mine
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Article Type:
Research Article
Full Title:
Comprehensive evaluation method of stope stability and its application in deep metal
mine
Short Title:
Comprehensive evaluation method of stope stability
Corresponding Author:
Hong Zhou
Northeastern University Key Laboratory of Safe Mining of Deep Metal Mines Ministry of
Education
Shenyang, CHINA
Keywords:
stope structural parameters; rock mass quality classification; stope stability; support
design
Abstract:
The problem of high stress condition is a critical challenge for mining activity in deep
metal mine, and the difficulty and economic cost of maintaining the stability of
surrounding rock also increase. In order to ensure the stability of surround rock mass
in the mining process, it is necessary to evaluate the stability of surround rock for a
more stable stope structure. However, there are many kinds of methods for the rock
stability assessment, most of them are based on the empirical and analytical methods,
which lead to an uncertain result. In this paper, a comprehensive evaluation method
including factors of stope stability, self-stabilization time, maximum safety span and
support measure is determined, this assessment method is more likely to provide a
comprehensive result. Finally, it was verified by the field application in Jiaojia Gold
Mine. Results show that it can accurately evaluate the stability of the surrounding rock
of the stope from different aspects compared with the traditional single evaluation
method, and it leads to a more comprehensive and reliable result.
Order of Authors:
Zhaojun Qi
Huanxin Liu
Zhihai An
Yantian Yin
Zhen Liu
Mingde Zhu
Yunsen Wang
Hong Zhou
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This work was funded by National Key R&D Programme of China (2021YFC3001304)
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Cover Letter
Dear Editors:
We would like to submit the enclosed manuscript entitled “Comprehensive evaluation
method of stope stability and its application in deep metal mine”, which is a research
article, and we wish to be considered for publication in “PLOS ONE”. No conflict of
interest exits in the submission of this manuscript, and manuscript is approved by all
authors for publication. I would like to declare on behalf of my co-authors that the work
described was original research that has not been published previously, and not under
consideration for publication elsewhere, in whole or in part. All the authors listed have
approved the manuscript that is enclosed.
In this work, we found the evaluation of stope stability is extremely important in the
process of metal mining, but there is no unified evaluation method at present, so we
proposed a comprehensive evaluation method according to the common evaluation
methods in engineering, which is more likely to provide a comprehensive result. In
addition, I have read some research articles related to rock mechanics and mining in
“PLOS ONE”, and hope this paper is also suitable for your journal.
We deeply appreciate your consideration of our manuscript, and we look forward to
receiving comments from the reviewers.
Thank you and best regards.
Yours sincerely,
Hong Zhou
E-mail: 2110404@stu.neu.edu.cn
Manuscipt
Click here to access/download;Manuscript;Manuscript.docx
1
Comprehensive evaluation method of stope stability and its application in deep metal
2
mine
3
Zhaojun Qi1, Huanxin Liu1, Zhihai An1, Yantian Yin1, Zhen Liu1, Mingde Zhu1, Yunsen Wang2,*,
4
Hong Zhou2
5
6
1 Deep
7
2
8
University, Shenyang, China.
Mining Laboratory of Shandong Gold Group Co., Ltd., Jinan, China
Key Laboratory of Ministry of Education on Safe Mining of Deep Metal Mines, Northeastern
9
10
*Corresponding
11
E-mail: wangyunsen@mail.neu.edu.cn(YW)
author
12
1
13
Abstract：
14
The problem of high stress condition is a critical challenge for mining activity in deep metal
15
mine, and the difficulty and economic cost of maintaining the stability of surrounding rock also
16
increase. In order to ensure the stability of surround rock mass in the mining process, it is
17
necessary to evaluate the stability of surround rock for a more stable stope structure. However,
18
there are many kinds of methods for the rock stability assessment, most of them are based on the
19
empirical and analytical methods, which lead to an uncertain result. In this paper, a comprehensive
20
evaluation method including factors of stope stability, self-stabilization time, maximum safety
21
span and support measure is determined, this assessment method is more likely to provide a
22
comprehensive result. Finally, it was verified by the field application in Jiaojia Gold Mine. Results
23
show that it can accurately evaluate the stability of the surrounding rock of the stope from
24
different aspects compared with the traditional single evaluation method, and it leads to a more
25
comprehensive and reliable result.
26
Key words: stope structural parameters; rock mass quality classification; stope stability; support
27
design
28
1 Introduction
29
Mining industry supports the sustainable development of the national economy. With the
30
reduction and depletion of shallow mineral resources, it has become an inevitable trend among the
31
world mining industry to utilizing deep resources. More than 80 metal mines, mostly in South
32
Africa, are over 1000 meters deep, such as TauTona Gold Mine in Witwatersrand (3910m), East
2
33
Rand Gold Mine (3585m), Mponeng Gold Mine (4500m), Anglogold Gold Mine (3700) and the
34
deepest platinum palladium mine in South Africa (2200m). Moreover, there are 16 metal mines
35
with a mining depth over 1000m in China. In the future, with the development of exploration
36
technology and equipment, it is entirely possible to find a large number of metal deposits in the
37
depth of 3000~5000m.
38
With the mining activity going deeper, the ground stress increases, the mining technical
39
conditions and environment deteriorate seriously, and the mining cost increases sharply. The
40
premise of mining is to ensure the stability of rock mass during the mining activity, so it is
41
necessary to design a reasonable stope structure parameter, which could ensure its stability. From
42
the actual production situation of mining site, the structural parameters of the underground stope
43
occupies a decisive role in its stability. Meanwhile, it affect the economic benefits of mining.
44
When this parameter is too large, it will lead to instability and caving of surround rock, which also
45
increase the dilution and risk of operation. When it goes too small, resulting in a large mining
46
volume, low ore recovery rate and economic efficiency(1, 2). Therefore, stope stability is an
47
important index to evaluate the rationality of stope structure parameters. Stope instability directly
48
affects ore loss and dilution, reduces mine economic benefits and even affects mine safety
49
production.
50
Many scholars have done lots of researches on stope stability. Mathews et al. First proposed
51
the stability diagram method in 1981(3), which is an empirical formula summarizing from a large
52
number of engineering examples. In 1988, Potvin revised Mathews stability chart by analyzing
53
242 cases and redefining some adjustment factors(4). In 1992, Potvin and Nickson et al. verified
54
the rationality of this method and proposed amendments by collecting more deep mining field data
3
55
(5). In 2000, Trueman et al. based on a large number of new examples of data, using the method of
56
logarithmic regression to redefine the stable area and serious damage area (6). Moreover,
57
Mawdeskey et al. gave the determined the equal probability diagram of these two areas in 2001
58
(7). The general step of using Mathews stability diagram method to evaluate the stope is to
59
calculate the stability number N’ and hydraulic radius HR on the basis of numerical simulation,
60
draw those parameters on the diagram, and evaluate stope stability according to the diagram area
61
where the intersection is located.
62
With the revolution of computer technology, the application of numerical analysis method of
63
rock mass engineering has become more extensive and necessary for studying the stability of
64
underground rock mass. Elrawy Abdellah et al.(8)used RS2D to study the depth of stope failure.
65
The stability of stope is evaluated according to the failure depth of plastic zone, the range of yield
66
or plastic zone and deformation profile.
67
Aiming at the problem that the goaf area is very dangerous, Wu et al.(9)used Hoek-Brown
68
strength criterion to establish the limit equation of pillar in this area under steady state and
69
introduced interval theory. The overlying strata weight, rock uniaxial compressive strength and
70
rock damage degree of interval variable value range as the condition to calculate the reliability and
71
then analyze the damage degree of rock mass, to evaluate the stability of goaf.
72
With the assistance of acoustic, optical or electrical sensors and equipment, the acoustic
73
signals of surrounding rock displacement, stress and rock fracture are collected and analyzed to
74
evaluate the stability of stope. At present, the main technology of goaf stability monitoring in
75
China are micro-seismic monitoring method, photo-elastic stress meter meter method, acoustic
76
emission monitoring method, three-dimensional laser scanning method, etc. In the process of
4
77
studying the stability of goaf in a gold mine, Zhao Kui et al.(10) used a long-term photo-elastic
78
stress meter to monitor the stress change in the mining activity, and carried out cluster analysis on
79
the monitoring results. On this basis, the stability of goaf is divided into three levels and pillars
80
that play a key role in it are identified.
81
However, there is no unified evaluation method for stope stability. According to the common
82
method in engineering, a comprehensive way is proposed in this paper. Firstly, the geomechanical
83
data is investigated, and then the RMR and Q values are calculated. Survey and calculation results
84
are input into the database. RMR, Q value and other information of the field point are calculated
85
by inverse distance weighted average method, Chinese national standard recommendation method,
86
self-stabilization time judgment method and Barton stope width empirical formula are used to
87
evaluate the stability of stope. At the same time, the support parameters are calculated by using the
88
support recommendation formula, Chinese national standard recommendation method and Q
89
grading support chart theory. These stability evaluation results and support parameters are
90
compiled into a table. Through the mutual confirmation of each result, a multi-dimensional
91
comprehensive evaluation system is formed from the aspects of stope stability, self-stabilization
92
time, safety span and support measures. In order to be quickly applied on site, the stability
93
evaluation system was developed by python and applied in Sizhuang mining area of Jiaojia gold
94
mine. Results show that the method and system provided in this paper can quickly evaluate the
95
stability of the designed stope, and the evaluation results are basically consistent with the actual
96
monitoring results.
5
97
2 Methodology
98
Through the investigation of rock mechanics parameters on site, using the results of rock
99
mass quality classification, combined with Mathews stability chart, limit span method for span,
100
Chinese national standard recommendation method, self-stability time judgment method and other
101
methods, the comprehensive evaluation of stope stability is carried out. At the same time, the
102
evaluation results are output into Excel table. This paper proposes a specific flow chart of
103
comprehensive evaluation method as shown in Fig 1.
104
105
Fig 1. Flow chart of comprehensive evaluation method
106
2.1 RMR and Q value calculation
107
Rock mass quality evaluation and classification are the basic work to evaluate the stability of
108
rock mass. Whereas the RMR and Q system are more familiar, and most of the stability evaluation
109
methods use index values of these two classification methods, so it is necessary to measure.
110
Rock mass rating is proposed by Bieniawski from CSRI of South Africa(11, 12). It consists
111
of five indexes: rock fragmentation, RQD, joint spacing, joint condition and groundwater. Each
112
value is scored according to the established criteria, and the RMR can be obtained by adding these
113
values.
114
𝑅𝑀𝑅 = 𝑅1 + 𝑅2 + 𝑅3 + 𝑅4 + 𝑅5 + 𝑅6
115
In the formula: R1 is the rock mass strength (point load strength or uniaxial compressive
116
strength); R2 stands for the core quality index RQD value; R3 represent for the joint spacing score
6
(1)
117
value; R4 is the joint condition score value; R5 is the groundwater condition value; R6 is the
118
modified parameter scoring system.
119
The Q system classification method was proposed by Barton of the Norwegian Institute of
120
Geotechnical Engineering in 1974(13). It is calculated by six indexes including rock mass quality
121
index, joint group number, roughness system, joint alteration coefficient, fissure water reduction
122
coefficient and in-situ stress conditions. Similar to the calculation of RMR, the above six indexes
123
are scored, and then the score value is substituted into the following formula to calculate the Q
124
value (14):
𝑄=
125
𝑅𝑄𝐷
𝐽𝑛
𝐽
𝐽
𝜔
∙ 𝐽𝑟 ∙ 𝑆𝐹𝑅
𝑎
(2)
126
Where 𝑅𝑄𝐷 is the rock mass quality index, 𝐽𝑛 is the number of joint group, 𝐽𝑟 is the
127
roughness system, 𝐽𝑎 is the joint alteration coefficient, 𝐽𝜔 is the fracture water reduction
128
coefficient, 𝑆𝑅𝐹 is the stress reduction coefficient.
129
2.2 Inverse Distance Weight Method
130
The investigation of rock mechanics parameters is a sampling survey in the stope or
131
development roadway. These survey data generally reflect the surrounding rock state of the
132
corresponding mining area, and cannot be specific to each point. The inverse distance weight
133
could be utilized to identifies the mechanical parameter in a specific point.
134
Inverse Distance Weight (IDW) is a spatially weighted average interpolation method can be
135
exact or smooth interpolation(15). It is used to calculate the attribute value of the unsampled point,
136
assuming that the value is a weighted average of the known values in the neighborhood(16). The
137
calculation method is as follows:
7
138
𝑃𝑖 = ∑𝑛𝑗=1 𝜆𝑗 𝑃𝑗
139
𝜆𝑖 =
(3)
𝑑𝑖𝑗 −𝛼
𝑛
∑𝑗=1 𝑑𝑖𝑗 −𝛼
(4)
140
∑𝑛𝑗=1 𝜆𝑗 = 1
(5)
141
Where 𝑃𝑖 is Parameter values of points to be solved; 𝑃𝑗 is Parameter values of known points in
142
the region 𝜆𝑖 is Weight of Known Points in Region 𝑛 is Number of known points in a region;
143
𝑑𝑖𝑗 is Distance from the point i to the point j; 𝛼 is Weight power function.
144
2.3 Stope Stability Calculation
145
2.3.1 Mathew Stability Chart Method
146
The Mathews stability diagram method proposed by Mathews et al.(17) is based on the
147
calculation of stability coefficient N and hydraulic radius S, and combines the stability diagram to
148
evaluate the stability of underground stope under supporting and non-supporting conditions. Then
149
Potvin(4) made improvements, and the expression of the stability coefficient is as follows :
𝑁=
150
𝑅𝑄𝐷
𝐽𝑛
∙
𝐽𝑟
𝐽𝑎
∙𝐴·𝐵∙𝐶
(6)
151
Where N is the stability coefficient; A is the rock stress coefficient; b is the correction coefficient
152
of joint orientation; c is the gravity adjustment coefficient of the designed exposed face; and other
153
symbols are the same as above.
154
Hydraulic radius HR reflects the size and shape of the stope. The hydraulic radius is the ratio
155
of the excavation surface to the perimeter of the exposed surface, which can be determined by the
156
following method shown in Fig 2, the hydraulic radius of the hanging wall is equal to the area
157
divided by perimeter.
8
158
Fig 2. Hydraulic radius determination method diagram
159
The calculation method of hydraulic radius S is :
𝑆=
160
161
162
163
𝐿×𝐻
2×(L+H)
(7)
Where L and H are the length and width of the exposed surface respectively.
After obtaining the stability coefficient and hydraulic radius, the stability of the stope under
the current structural parameters is judged according to Fig 3.
164
165
Fig 3. Mathews stability chart
166
2.3.2 Limit Span Method
167
The research team of UBC University in Canada obtained the relationship between rock mass
168
quality RMR and the limit span of the stope by analyzing 292 examples of stope roof stability
169
situation mined by different layered filling methods in Canada(18), and established the limit span
170
curve, as shown in Fig 4(19). It can be seen from Figure 2.3 that the limit span curve divides the
171
surveyed observation points into three regions: stable, potential unstable (fluctuation) and unstable.
172
Based on this, the stope span can be designed or the stability of the surrounding rock under a
173
certain span can be evaluated. It should be noted that the limit span in this method is defined as the
174
diameter of the maximum inscribed circle within the exposed roof boundary, and its determination
175
method is shown in Fig 5.
176
177
Fig 4. Limit Span Method (Under the condition of no support or local support)
178
Fig 5. Determination of limit span of stope
9
179
In addition, since the limit span method is based on the RMR evaluation method, when there
180
are obvious gently inclined joints (joint dip angle is less than 30 degree), the RMR value in this
181
area needs to be further reduced by 10.
182
2.4 Stope self-stabilization time calculation
183
2.4.1 Self-stabilizing time evaluation (based on RMR classification)
184
The stability time of surrounding rock is also an important index to evaluate the stability of
185
surrounding rock. Its value can not only reflect the quality of surrounding rock in stope, but also
186
provide suggestions for the selection of excavation and support method. In the study of
187
self-stabilizing time, Bieniawski(20) summarized the relationship between the structural
188
parameters of the unsupported stope (exposed surface span) and its self-stabilizing time in 1993
189
based on the RMR classification of rocks, as shown in Fig 6 The self-stabilizing time can be
190
estimated according to the RMR value of the surrounding rock and the unsupported span, and the
191
larger the RMR value and the smaller the span design, the longer the self-stabilizing time of the
192
surrounding rock.
193
Fig 6. Self-stabilization time judgment based on RMR classification
194
2.4.2 Chinese national standard recommendation method
195
The above self-stabilizing time assessment method only obtains the self-stabilizing time
196
according to RMR and span, which result is slightly rough. Based on this, China’s national
197
standard “Technical Code for Rock and Soil Anchor and Shotcrete Support Engineering” (GB
10
198
50086-2015)(21) determined the provisions on the level of surrounding rock of tunnel caverns,
199
which makes the evaluation from the aspects of rock mass structure, rock strength, rock mass
200
acoustic wave and rock mass stress ratio, and uses multiple indicators to judge the stability of
201
stope or tunnel without support. Those results not only estimate the self-stabilization time, but also
202
give the failure form of surrounding rock.
203
Rock mass integrity index Kv can be calculated as follows :
𝐾𝑣 = (
204
𝑉𝑝𝑚 2
)
𝑉𝑝𝑟
(8)
205
Where 𝑉𝑝𝑚 is Measured Longitudinal Wave Intensity of Tunnel Rock (km/s); 𝑉𝑝𝑟 is Measured
206
longitudinal wave velocity of tunnel rock (km/s).
207
208
209
210
211
When the acoustic wave is measured unconditionally, the volumetric joint count Jv of rock
mass can also be used to determine the Kv value according to Table 3.2.
Rock mass strength stress ratio in surrounding rock classification table should be calculated
according to the following formula:
(1) When there is a measured number of ground stress :
𝐾𝑣 𝑓𝑟
212
𝑆𝑚 =
213
Where 𝑆𝑚 is Rock mass strength stress ratio; 𝑓𝑟 is Saturated uniaxial compression strength of
214
rock (kPa); 𝐾𝑣 is Integrity coefficient of rock mass; 𝜎1 is The maximum principal stress of
215
vertical hole axis (kN/m2).
216
217
218
219
𝜎1
(9)
(2) When there is no in-situ stress measured data, 𝜎1 can be determined by the following
formula or by displacement back analysis data:
𝜎1 = 𝛾𝐻
(10)
Where 𝛾 is Gravity density of rock mass (kN/m3); 𝐻 is Thickness of tunnel roof overburden
11
220
(m).
221
2.5 Calculation of Maximum Safety Span of Stope
222
The reasonable design span is an important factor to ensure the stability of the stope.
223
Therefore, determining the maximum safety span of the stope is also one of the indexes to
224
evaluate the stability of the stope. For the calculation of stope safety span, Barton et al.(22)
225
suggested the following empirical formula to determine the engineering span through field
226
investigation
227
𝑊 = 2 ∙ 𝐸𝑆𝑅 ∙ 𝑄 0.4
228
Where W is the maximum safety span of no support system, Q is rock mass quality index, ESR is
229
the support ratio, which is generally valued according to Table 1. In mining activity, the ESR value
230
is generally selected as 3 according to experience.
231
(11)
Table 1. ESR value
Project Type
A. Temporary engineering
ESR
2~5
B. Permanent mining roadways, water tunnels, auxiliary tunnels, roadways and
1.6~2
chambers
C. Storage chambers, water treatment works Small road and railway tunnels
1.2~1.3
D. Power stations, main road and railway tunnels, etc.
0.9~1.1
E. Underground nuclear engineering base, station, public place, factory or major gas
0.5~0.8
supply tunnel
12
232
2.6 Stope Support Parameter Calculation
233
The calculation of stope support parameters is also an important content to evaluate the
234
stability of stope and roadway. On the one hand, it provides reasonable suggestions for stope
235
support, on the other hand, it can check the rationality of the supported stope. This comprehensive
236
evaluation method uses empirical formula, national standard recommended method and Q grading
237
support chart theory to calculate support parameters.
238
2.6.1 Empirical formula (bolt spacing and length)
239
According to the existing experience and examples of bolting and shotcreting support, the
240
following empirical formula can be used to determine the bolt parameters for mining projects with
241
a span of less than 10 m :
242
(1) Bolt length L:
𝐿 = 𝑛(1.1 +
243
𝐿 > 2𝑆
244
𝐵
10
)
(12)
(13)
245
Where B is the roadway span. N is the stability parameters of surrounding rock. For the
246
surrounding rock with good stability, N is 1.0, for the surrounding rock with poor stability, N is
247
1.1, for unstable surrounding rock, N is 1.2. S is the joint spacing in surrounding rock. L is the
248
length of bolt. Bolt length takes the larger value according to the above calculation. In domestic
249
roadway engineering, the length of bolt is generally from 1.5m to 2.0m, whereas 2m to 3m in
250
stope and 2m to 5m in large chamber engineering.
251
(2) Bolt spacing D
13
252
𝐷 < 0.5𝐿
(14)
253
𝐷 > 3𝑆
(15)
254
The fracture spacing S is an important factor to determine the bolt spacing. Domestic
255
roadway engineering, bolt spacing is generally 0.8m to 1.0m, the longest not more than 1.5m.
256
(3) Bolt diameter d
257
Bolt diameter depends mainly on the type of anchor and anchorage force：
𝑑≈
258
𝐿
110
(16)
259
If the bolt length is 2200 mm, the diameter d equals to 20 mm.
260
However, the above empirical formulas and data ignore the anchoring force of the bolt. The
261
parameters selected by empirical formula, especially the anchorage force, must be tested by
262
theoretical calculation.
263
2.6.2 China national standard recommended method (anchor spray
264
support)
265
China national standard “Technical code for engineering of ground anchorages and shotcrete
266
support” (GB 50086-2015)(21) provides a “Design parameter chart of anchor spray support in
267
slant hole and tunnel”. It demonstrates detailed suggestions for the selection of support parameters
268
of excavation cavern in combination with surrounding rock grade and excavation span. In the
269
process of selecting support parameters, the type and parameters of bolting and shotcrete support
270
are initially selected according to this chart, and then the mining parameters are modified
271
according to the actual engineering. In addition, when the cavern excavation span is greater than
272
20m and the height-span ratio H/B is greater than 1.2, the sidewall supporting parameters should
14
273
be strengthened according to the specific circumstances of the project. When the height-span ratio
274
H/B of the cavern is greater than 2.0, the sidewall support should be strengthened by pre-stressed
275
anchor group support with a length not less than 0.3 times the height of the side wall. The rock
276
pillars between caverns are strengthened according to their thickness or supported by through-type
277
pre-stressed anchor bolts.
278
2.6.3 Q classification support chart theory
279
The Barton rock mass quality classification Q system links the rock mass quality with the
280
type of support, and the recommendation of support type is very detailed. According to the Q
281
value, the support design can be divided into 9 categories, as shown in Figure 7(23). It shows the
282
latest classification and support charts (based on the statistical results of 2000 tunnels) in the
283
“Norwegian Tunnel and Underground Engineering 2004 Annual Report” published in 2004. It is a
284
series of De-Q value curves from which the required support types can be determined.
285
286
287
The De in the figure is the equivalent size, which is the ratio of the engineering span to the
excavation support ratio. The calculation formula is as follows :
𝐷𝑒 =
Excavation space span
ESR
(17)
288
289
Fig 7. Q system rock classification and support chart
290
2.7 Comprehensive Evaluation of Stope Stability
291
The information of rock mechanics parameters investigated is input into a special database.
292
At the same time, in order to better and faster realize the field application, a comprehensive
15
293
evaluation system of stope stability is compiled by python to realize the comprehensive evaluation
294
of the above various evaluation methods. For the graph curve involved, getdata and origin are
295
used to get the curve information and fit it.
296
In the process of evaluation, the information such as the middle section name, coordinate
297
value and mining parameters of the point that needs to be judged for stability is input to the
298
comprehensive evaluation system of stope stability. The system will find the database to find the
299
parameters of the same middle section according to the input parameters, and then use the inverse
300
distance weight difference method to calculate the parameter information of the input point, so as
301
to evaluate the stability, and output the evaluation results of multiple methods to a result report.
302
The generated result report shows the stability evaluation results of four aspects: surrounding rock
303
stability, self-stabilization time of exposed surface, maximum safety span, support parameters and
304
measures.
305
(1) Stability evaluation:
306
The Mathews stability chart method and the UBC limit span method are used to evaluate the
307
stability of the input mining parameters at the current evaluation point. The Mathews stability
308
chart method gives three evaluation levels: stability, destruction, and caving. The UBC limit span
309
method gives three evaluation levels: stability, fluctuation, and instability. The two evaluation
310
results confirm each other and get more accurate stability evaluation.
311
(2) Self-stabilizing time evaluation :
312
The self-stabilization time judgment method based on RMR classification obtains its span
313
through the output mining parameters, and then combines the RMR value of the mining point to
314
determine the self-stabilization time of the surrounding rock ( in hours ), while the Chinese
16
315
national standard recommendation method is based on the self-stabilization time evaluation from
316
the aspects of rock mass structure, rock strength, rock mass acoustic wave, rock mass stress ratio,
317
etc., and outputs the self-stabilization time and failure form of the mining point in the span of
318
5-10m or below 5m. The two judgment results can confirm each other. If the difference is large, it
319
is necessary to check the rationality of the data in the database and re-evaluate. If the difference is
320
small, it is considered that the judgment result is reasonable, and the smaller value is taken as the
321
self-stabilizing time evaluation value of the evaluation point.
322
(3) Maximum safety span
323
The stability evaluation system obtains the Q value of the judgment point in the database,
324
takes ESR value as 3, and then uses the empirical formula to calculate and output the maximum
325
safety span of the judgment point. In addition, combined with the evaluation results of the UBC
326
limit span method, the rationality of the evaluation is tested. If the two values differ greatly, it is
327
necessary to check the validity of the data. If the difference is not large, a smaller value can be
328
taken as the evaluation result of the maximum safety span.
329
(4) Support parameter calculation
330
The empirical formula gives the bolt support suggestion of the judgment point under the
331
input mining parameters, the national standard recommendation method gives the bolt-shotcrete
332
support type and parameter suggestion, and the Q classification support chart theory gives the
333
support parameter design suggestion. The specific implementation of supporting measures has
334
certain subjectivity. The above evaluation methods provide recommended values for the design or
335
optimization of supporting parameters of stope or roadway. The final supporting parameters need
336
to be corrected by mine workers or designers according to the field situation.
17
337
3 Field application example
338
3.1 Engineering background overview
339
Jiaojia Gold Mine is the core enterprise of Shandong Gold Mining Co., Ltd. It has three
340
mining areas: Jiaojia, Wangershan and Sizhuang mining area. Among them, the ore-controlling
341
structure of Jiaojia mining area is the main fault of Jiaojia type in the hanging wall, and the ore
342
body occurs in the footwall alteration zone of the main fault. The overall trend of ore body is
343
NE54°, tendency NW, dip angle of about 27°. The horizontal thickness of the ore body is about
344
2m to 70m, the thickest is about 70m near the 104 line, and the two wings gradually become
345
thinner to less than 10m. The ore body belongs to gold-bearing pyritization, yellow iron sericite
346
quartzite broken altered rock type, the hanging wall of ore body and the surrounding rock are fault
347
contact relationship, the boundary is obvious ; the footwall of the ore body has a gradual transition
348
relationship with the surrounding rock, and there is no obvious boundary. The hanging wall rock
349
of ore body is amphibolite with poor stability, large exposed area or long residence time, which is
350
prone to collapse. The footwall of the ore body is sericitization, silicification and potassic granite,
351
and there are also interlaced fracture joints in it, resulting in local surrounding rock fragmentation.
352
The main mining methods include downward drift cemented filling mining method, upward
353
horizontal drift cemented filling mining method and medium-deep hole double-amplitude
354
sub-section subsequent filling method.
355
356
357
Fig 8. Location of Jiaojia Gold Mine
At present, the mining method of Jiaojia Gold Mine mainly has the following problems :
18
358
(1) The mining method of upward horizontal entry cemented filling has been used for many
359
years in the middle production stage of Jiaojia Gold Mine. The entry section size of most stopes is
360
controlled to 3.5m × 3.5m. The small entry section is beneficial to the safe mining operation, but
361
there are also some problems such as small stope size, small production capacity of single stope,
362
low production efficiency, high production cost and complex construction organization. The
363
production task of Jiaojia Gold Mine in 2021 is 7.6 tons of gold. The mine faces great pressure to
364
successfully complete the production task. It is difficult to successfully complete the heavy
365
production task by continuing to adopt the small-size mining design.
366
(2) The current supporting means of Jiaojia Gold Mine mainly include : pipe seam anchor
367
support, anchor net support, anchor net spray support and U-shaped steel arch support. Especially
368
in the stope, the roof joint fissure development, part of the existence of two groups or even
369
multiple groups of joint staggered distribution phenomenon, only through the 1.8 m pipe joint bolt
370
and through the belt to support, often can not meet the requirements of maintaining roof safety.
371
In summary, it is urgent to count the distribution of joints and fissures in Jiaojia gold mine,
372
classify the stability of ore bodies exposed in different stopes, and optimize the industrial test of
373
upward horizontal approach filling method parameters according to the classification results of ore
374
rock stability, so as to realize efficient mining and successfully complete the production
375
connection task of Jiaojia mining area.
376
This research mainly aims at the rock mechanics data collection of -570m main production
377
middle section of the mine, and carries on the experimental design stope stability appraisal
378
according to the rock mechanics data and the mine reality.
379
19
380
Fig 9. Experimental evaluation of stope design
381
3.2 Rock Mechanics Parameters Acquisition
382
3.2.1Investigation of Rock Mass Structural Plane
383
In practical engineering, these small and numerous structural planes are likely to cause
384
large-scale rock engineering problems, that is, rock mass instability. Therefore, in order to explore
385
the failure mechanism of the surrounding rock of the roadway with more developed joints, it is
386
necessary to analyze the characteristics of joint development, such as distribution law, surface
387
characteristics and spatial combination form. In this study, the Sirovision rock mass telemetry and
388
structural analysis system was selected to scan the structural planes of the three-layer and
389
four-layer approach stopes of a sub-lane in the level of-570 m in the Sizhuang mining area of
390
Jiaojia Gold Mine, and the development of joints and fissures in the area was obtained by software
391
splicing. Combined with Dips6.0 software, the joint measurement results are grouped and counted,
392
and the joint Schmidt diagram and joint rose diagram are drawn. The results are shown in Fig 11.
393
The information of -570m level approach structural plane in Sizhuang mining area of Jiaojia Gold
394
Mine is summarized in Table 3. Based on this, Q system and RMR rock mass quality evaluation
395
system are used to classify the rock quality of the surrounding rock in the above area, which
396
provides data basis for the optimization of roadway support and the selection of stope span
397
parameters in this area.
398
Fig 10. Drawing of rock mass structural plane of measuring point in -570m level by
399
Sirovision
20
400
Fig 11. Schmitt diagram and rose diagram of joint in -570m level (a) Schmitt diagram; (b) rose
401
diagram
402
Table 2. Structural plane information in -570m level
Joint
Joint
Location
Dip
Group
Number
(o)
Joint
Volume
Spacing
Joint
（cm）
Number
Dip Angle
Condition of Joint Structural Plane
(o)
Number
One lane
1
302.7
55.2
16
Three layers
The joint fissures of rock mass are
2
11.08
-570m level
relatively developed, and the joints
2
89.2
80.3
74
(9360,9120,-570)
are mostly wavy or rough and
One lane
irregular. There are fillers between
1
308.5
59.6
22
some joints, and the structural plane
Four layers
2
20.93
is relatively wet.
-570m level
2
90.7
53.0
27
(9350,9130,-570)
403
3.2.2 Rock Mechanics Experiment
404
The strength of rock mass has engineering guiding significance for the implementation of ore
405
mining. However, some basic physical and mechanical parameters of rock are needed to calculate
406
the strength of rock mass, so these parameters are the basis of the calculation of the whole stress
407
field. At the same time, the different properties of rock under different conditions have certain
408
reference value for the methods and strategies of field construction.
409
The sources of the indoor test samples are mainly roadway sampling and stope sampling. The
21
410
test rock is mainly granite rock samples. The ZTR-276 rock triaxial test machine is used, and the
411
maximum load is 2000 kN. This includes a loading system, a test system and a control system, as
412
shown in Fig 12. The device has different control modes such as axial stress control, axial stroke
413
control, axial strain control and cyclic strain control, which can meet the requirements of various
414
experimental designs.
415
416
Fig 12. ZTR-276 rock triaxial test system
417
During the experiment, the loading is controlled by longitudinal strain, and the loading
418
control rate is (1~5)×10-6mm/s. The computer automatically records the original data such as
419
displacement and load of the experiment, and then the stress-strain relationship curve can be
420
drawn after secondary processing. Fig 13 shows the whole process curve of loading and unloading
421
stress-strain of some samples in the middle section of -570m.
422
423
Fig 13. Process curve of rock uniaxial compression in -570m level of Sizhuang Mining Area
424
The uniaxial loading experiment was carried out on the rock in -570m level of Sizhuang
425
Mining Area, and the stress-strain curve was drawn by the experimental data. It can be seen from
426
the figure that the early curve of rock compressive strength is not smooth, indicating that the rock
427
has joint cracks. During the loading process, the joint cracks weaken the bearing capacity of the
428
rock, and the cement in the rock joint cracks affects the bearing capacity of the rock. There is a
429
small peak curve before the peak strength of the stress-strain curve of the rock sample SZY-008,
430
indicating that there are joints in the rock sample, and its failure strength cannot truly reflect the
431
compressive strength of the rock itself. Therefore, the compressive strength of the rock is 140
22
432
MPa. There are many post-peak curves of SZY-012 sample, and it has obvious characteristic
433
curves. After the sample passes the peak stress, its internal structure is destroyed, and the single
434
sample basically maintains the whole shape. In the post-failure stage, the fractures develop rapidly,
435
intersect and combine with each other to form macroscopic fracture surfaces. Since then, the
436
deformation of the rock sample is mainly manifested as the block slip along the macroscopic
437
fracture surface. The bearing capacity of the sample decreases sharply with the increase of
438
deformation, but it does not drop to zero, indicating that the rock sample still has a certain bearing
439
capacity after fracture, but maintains a small value. The strength relative to the final point is called
440
residual strength. With the increase of load until failure, the volume of rock sample increases
441
rather than decreases. This phenomenon is called dilation. The failure of rock sample undergoes a
442
dilation stage, so dilation is often a precursor of rock failure. The results of uniaxial compression
443
test are shown in Table 3.
444
Table 3. Uniaxial compression test results of -570m level rock
Sample Size (mm)
Average
Compressive
Compressive
Location
Lithology
Strength
Diameter
Height
Strength
(Mpa)
(Mpa)
One lane
Three layers
Sericite
-570m level
Rock
(9360,9120,-570)
One lane
Sericite
Four layers
granite
49.80
99.80
63.07
49.60
99.60
136.69
49.70
99.90
129.42
49.60
99.66
144.75
133.05
134.10
49.62
99.75
23
120.27
-570m level
49.76
100.00
137.28
(9350,9130,-570)
445
3.3 Calculation of RMR and Q Value
446
Through the investigation of the surrounding rock in the -570m level in Sizhuang mining area of
447
Jiaojia Gold Mine and the rock mechanics experiment, the rock quality classification of the
448
surrounding rock in the above area is carried out by using Q system and RMR evaluation system,
449
and the classification results are shown in Table 4.
450
Table 4. Rock classification summary
Q
RMR
Location
Value
Description
Value
Grade
Description
29.42
Good
71
Ⅱ
Good
8.61
General
55
III
General
One lane
Three layers
-570m level
(9360,9120,-570)
One lane
Four layers
-570m level
（9350,9130,-570）
451
3.4 Stability Evaluation
452
The investigated rock mechanics parameter information and rock quality classification results
453
will be input into the rock mechanics parameter database associated with the comprehensive
454
evaluation system of the stope. After inputting the location information and mining parameters of
24
455
the proposed optimization stope or roadway, the system will extract and calculate the
456
corresponding parameters, and evaluate the stability of the area where the point is located. At the
457
same time, it provides suggestions for support optimization and stope span parameter selection.
458
In this study, the point (9360,9120,-570) at the three-layer mining and the four-layer
459
(9350,9130,-570) in the Sizhuang mining area of Jiaojia Gold Mine were selected for evaluation.
460
Among them, the stope in three-layer is 59.84 m long, 5.16m wide and 4.01m high, compared
461
with the stope in four-layer is 31.87m long, 3.5m wide and 3.49m high. Input these information
462
into the stope comprehensive evaluation system, the system will automatically output the results
463
report in the form of excel after calculation.
464
3.5 Stability Evaluation Results and Discussion
465
466
For the evaluation points selected in Section 3.4, the main evaluation results of the output
report of the stope comprehensive evaluation system are shown in Table 5 and Table 6.
467
Table 5. Results of three-layer
Stope
Jiao Jia Gold Mine, -570m level, One lane, Three layer
Coordinate
（9360,9120,-570）
Proposed design
Stope
parameters
Length: 8m, Width: 5m, Height: 4m
Rock Classification
RMR
71
Q
29.42
Stope stability
UBC limit span method
Scheme in stable area, limit span is 26.95 meters
evaluation
Mathews stability chart method
When the span is 5m to 10m, rock can be stable
China national standard
Self-stabilizing
for a long time (months to years), only a small part
recommendation method
of it falls off
time judgment
Self-stabilizing time judgment
Maximum safety
span without
Scheme in stable area
Barton stope width empirical formula
25
The self-stabilizing time without support after
mining exposure is estimated to be 42137.59 hours
The maximum safety span of unsupported area is
23.21m
support
Bolt support length 1.76m,
Empirical formula
Bolt spacing less than 0.88m,
Bolt diameter 16m
Support measures
and parameters
China national standard
Shotcreteδ=50
recommendation method
Q classification support chart theory
468
No support required
Table 6. Results of four-layer
Stope
Jiao Jia Gold Mine, -570m level, One lane, Four layer
Coordinate
（9350,9130,-570）
Proposed design
Stope
parameters
Length: 8m, Width: 5m, Height: 4m
Rock
UBC limit span method
55
Classification
Mathews stability chart method
8.61
Stope stability
evaluation
Self-stabilizing
time judgment
Maximum safety
span without
support
China national standard
Scheme in stable area, limit span is 16.33 meters
recommendation method
Self-stabilizing time judgment
Scheme in stable area
Barton stope width empirical
When the span is 5m, the stability time of surrounding
formula
Empirical formula
China national standard
recommendation method
rock is very short, about hours to days
Self-stabilizing time without support after mining
exposure is estimated to be 2454.09 hours
Maximum safety span of non-supporting area is 14.20m
Bolt support length 1.74m,
Q classification support chart
Bolt spacing less than 0.87m,
theory
Bolt diameter 15.42m
Support measures
and parameters
1. Shotcreteδ=80~100
UBC limit span method
2. Shotcreteδ=50, Bolt length 1.5~2.0m, @0.75~1.0
Mathews stability chart method
No support required
469
From Table 5, it can be seen that the surrounding rock quality grade of the three-layer is good.
470
Under the proposed mining parameters, the Mathews chart method and the UBC limit span
471
method are evaluated in the stable area, and the stope is in a stable state. At the same time, under
472
the proposed mining parameters, the national standard recommended method predicts that the
473
surrounding rock of the stope can maintain stability for a long time (months to years), which is
474
also consistent with the 42137.59 hours (about 4 years and 10 months) given by the
26
475
self-stabilization time judgment. In addition, the maximum safety span of the unsupported area in
476
the Barton stope width empirical formula is 23.21m, and the limit span is 26.95m in the UBC limit
477
span method, which is similar to the Barton stope width empirical formula. Therefore, the
478
maximum safety span of the unsupported area is 23.21m, and the proposed stope structure
479
parameters are within a reasonable range. Finally, the comprehensive evaluation system from the
480
empirical formula, the national standard recommended method, Q classification support chart
481
theory three methods are given different support measures support parameters, which need to be
482
combined with the actual selection of support methods for reference.
483
From Table 6, it can be seen that the surrounding rock quality of four layers is general. Under
484
the proposed mining parameters, Mathews chart method and UBC limit span method are used to
485
evaluate it in the stable area, and it is concluded that the stope is in a stable state. At the same time,
486
under the proposed mining parameters, the national standard recommended method points out that
487
when the span of the cave is 5m, the stability time of the surrounding rock is very short (about
488
several hours to several days), which is different from the 2454.09 hours (about 3 months) judged
489
by the self-stabilization time method. On the one hand, because the national standard method
490
gives the stability of the 5 m span of the cave, and the evaluation of the stope span is 3.5 m, on the
491
other hand, because the national standard method is located in the standard " Technical
492
Specification for Geotechnical Bolt and Shotcrete Support Engineering " proposed by the Ministry
493
of Housing and Urban-Rural Construction of the People 's Republic of China. The application in
494
mines still has certain limitations, but no matter which method. Both reflect the poor stability of
495
the surrounding rock of the stope, and the design of reducing the stope span or strengthening the
496
support can be considered. In addition, the maximum safety span of the unsupported project in the
27
497
Barton stope width empirical formula is 14.20m, and the limit span is 16.33m in the UBC limit
498
span method, which is similar to the Barton stope width empirical formula. Therefore, the
499
maximum safety span of the unsupported project is 14.20m, and the proposed mining parameters
500
are within a reasonable range. Finally, the comprehensive evaluation system proposes the support
501
parameters of different support measures from the three methods of empirical formula, national
502
standard recommendation method and Q classification support chart theory, which needs to be
503
further judged in combination with the actual support method selected on site.
504
3.6 Actual Parameters and Effect of Stope
505
In order to verify the accuracy of the system in judging the stability of the stope, the stability
506
of both layer was investigated, as shown in Fig 14. The main production section of Jiaojia Gold
507
Mine has been mined by upward horizontal approach cemented filling mining method for many
508
years. The existing support methods mainly include bolt through belt support, shotcrete support,
509
U-shaped steel arch support, anchor net support and combined support of various situations.
510
511
Fig 14. Stope investigation (a)Stope investigation; (b)Work face of three layer; (c)Work face of
512
four layer
513
According to the previous rock mechanics investigation, surround rock in three layer belongs
514
to grade III good rock mass. The final acceptance parameters of the stope are 59.84m long, 5.16m
515
wide and 4.01m high. The support is carried out in the form of 1.8m bolt support, the anchorage
516
spacing is 1m and the row spacing is 3m. The investigation of the ground pressure behavior of the
517
stope for 40 days shows that the stope is relatively stable within 40 days, and there is no obvious
28
518
ground pressure behavior. According to the field survey results, the stope is in a stable state, the
519
acceptance span is much smaller than the maximum safety span of the unsupported area of 14.20
520
m, and the self-stability time is longer, which is consistent with the evaluation results of 3.5
521
sections, and the stope structure parameters are reasonable. At the same time, the stope using bolt
522
support, the parameter selection is better than the recommended parameters in Table 3.5, no
523
optimization is required.
524
Rock mass in four layer belongs to II level general rock. The final acceptance parameters of
525
the stope are 31.87m long, 3.5m wide and 3.49m high. The form of bolt-shotcrete combined
526
support is adopted, in which the thickness of shotcrete is 50mm, the length of bolt is 1.8m, the
527
spacing of bolt is 1m, and the row spacing of bolt is 3m. In the investigation of the ground
528
pressure behavior of the stope for 14 days, it can be seen that there was a spalling on the 6th day
529
after excavation and a falling block on the 13th day. From the survey results, it can be seen that
530
although the comprehensive evaluation method gives that the stope is in a stable state under the
531
proposed mining parameters (Table 3.6), and the acceptance span is much smaller than the
532
maximum safety span of the unsupported area 14.20m, but the actual excavation situation is not
533
optimistic, that is, on the 6 th day after excavation, the situation of spalling and falling occurs, and
534
in fact, this instability situation is also roughly consistent with the judgment result of the
535
self-stability time in the comprehensive evaluation method. It can also be seen that the stability
536
evaluation of the surrounding rock of the stope is multifaceted. The judgment is relatively simple
537
from the span and rock quality classification, while the comprehensive evaluation method adds
538
more evaluation factors. Using multiple evaluation methods to evaluate from multiple dimensions
539
can not only self-test the rationality of each result, but also avoid the limitations of a single
29
540
evaluation method to a certain extent. In addition, in terms of support, the stope adopts
541
bolt-shotcrete combined support, and its support parameters are consistent with the recommended
542
values in Fig 15. However, according to the results of self-stabilization time judgment and field
543
investigation, it needs to be further strengthened.
544
545
Fig 15. Stope failure in four layer (a) Bolt support failure; (b) Roof Caving; (c) Cracking of side
546
concrete; (d) Roof concrete caving
547
4 Conclusion
548
In this study, a comprehensive evaluation method of stope stability is proposed. This method
549
combines Mathews stability chart, limit span method, Chinese national standard recommendation
550
method, self-stabilization time judgment method, Barton stope width empirical formula and Q
551
grading support chart theory. The stability of surrounding rock of stope or roadway is evaluated
552
from four aspects : surrounding rock stability, self-stabilization time, safety span and support
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parameters, and reasonable suggestions are put forward for the selection of support parameters and
554
measures. The test was carried out in the Sizhuang mining area of Jiaojia Gold Mine. The results
555
show that the comprehensive evaluation method can accurately evaluate the stability of the stope
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or surrounding rock, and compared with the traditional evaluation method, the evaluation results
557
are more abundant. The sub-evaluation methods can confirm and compare each other, and then
558
provide a more comprehensive and accurate reference for mine design.
559
However, it still has limitations and imperfections at present: (1) In order to ensure the
560
multi-dimensional evaluation results, the comprehensive evaluation method uses two non-mine
30
561
field surrounding rock evaluation methods, namely the Chinese national standard recommendation
562
method mentioned in 2.4 and 2.6 sections, which may interfere with the verification and
563
inspection of the evaluation results. (2) The comprehensive evaluation method pays attention to
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the multidimensional nature of the evaluation, only simply integrates a variety of surrounding rock
565
stability evaluation methods, the output results are more dispersed, and no specific suggestions are
566
put forward for the optimization of structural parameters or support parameters. In view of these
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shortcomings, in the future, this study is committed to the continuous enrichment and
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improvement of the comprehensive evaluation method, try to organic integration of various
569
evaluation methods, and combined with the economic benefit analysis of the mine, output more
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specific and intuitive structural parameters and support parameters optimization value in the
571
results report. In addition, it is necessary to optimize the implementation means of field
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application, such as the development of evaluation system suitable for mobile communication
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equipment based on comprehensive evaluation method, which is convenient for mine workers and
574
designers to evaluate stope stability dynamically and quickly.
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Acknowledgment
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This work was funded by National Key R&D Program of China (2021YFC3001304) and the
577
Fundamental Research Funds for the Central Universities(N2101044).
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Conflicts of Interest
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The authors declare that the research was conducted in the absence of any commercial or
financial relationships that could be construed as a potential conflict of interest.
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