Drill, blast practices to improve blasting efficiency at Murgul copper mine

Category Mine
Group GSI.IR
Location 20th WORLD MINING CONGRESS 2005
Author Turker Hudaverdi Cengiz Kuzu
Holding Date 14 January 2006
 
ABSTRACT
 
The paper focuses on the influence of blast geometry and initation pattern for blasting optimization. Murgul Open-pit Copper Mine located in north-eastern Turkey was chosen for field studies. Firstly, in this study, geological and geotechnical parameters of working site were investigated and the production technique of Murgul Copper Mine was analysed. Drilling operation was observed and the effect of the drilling efficiency on multi-row blasting was evaluated. Blasting technique was analysed and appropriate blasting design parameters had been investigated using the approaches of blasting researchers. As a result of on-site observations as well as analytical studies, four different blasting patterns were determined and experimental blasts were performed in order to improve drilling-blasting works. On-site results of experimental studies were analysed and assessed by using proposed performance analysis methods (boulder counting, excavator loading performance analysis, particle size distribution analysis). There are also drilling-blasting cost analysis for current drilling-blasting design and experimental cases. Because of the centre of population around the Murgul Copper Mine; environmental effects of the blasting activities were evaluated. Blasting induced vibrations and flyrock were examined. As a result of economical, technical, and environmental outcomes, the most appropriate pattern was suggested as a solution for Murgul Copper Mine.
Key words: Drilling pattern, Blasting, Vibrations, Flyrock,
 
 

INTRODUCTION
 
Drilling and blasting is dominant operation for excavation of any rock, ore and minerals. Inappropriate drilling pattern causes high operational costs and loss of productivity during excavation, loading and transportation operations. Applying optimum drilling and blasting pattern is the first key to work productively and profitably.
 
So many parameters affect the design of drilling and blasting. While some of these parameters can be controlled, others cannot be. Controllable parameters are hole diameter, hole length, burden, hole inclination, stemming, dimension of blasting group, explosive type, ignition system, etc. Uncontrollable parameters are rock features and mass characteristics such as density, compressive strength, tensile strength and joints. In design stage, all parameters should be considered and evaluated together to obtain optimum working conditions.
 
The unit cost of drilling and blasting averages about 56% of total unit cost of mining in Murgul open pit mine. Primary blasting operation produces poor rock fragmentation leading to a requirement for secondary breakage and this increases costs. Improving the design of the drilling and blasting pattern can reduce drilling blasting costs.
 
The current operational and geological conditions are first found out and summarized based on the field observations, in this study. The effect of the drilling efficiency on blasting operations is investigated. A set of new drilling patterns for Murgul is searched based on the empirical equations developed by previous researchers. Experimental blasts are performed by using these drilling patterns, keeping the other blasting parameters constant. Then, the efficiency of the blasts is evaluated by visual observations, size distribution of fragmented rock particles, boulder counting, shovel loading rate and cost analysis. Blast induced vibration and flyrock problems are investigated to evaluate environmental aspect of the blasting operations.
 
TEST SITE DESCRIPTION
 
The test site is located close to Murgul located about 48 km away from the city Artvin in North-eastern Turkey. Murgul open pit copper mine that is a subsidiary of Black Sea Copper Works (KBI) has been operated since 1973. Current production rate of the mine is about 2,500,000 tons/year.
 
Concentrate copper processed in Murgul plant is about 75,000 tons/year. Concentrate copper is transported by two pipelines to the town Hopa located on the cost of Black Sea. After filter pressing and drying, the concentrate is transported to the smelting plant in the city Samsun by ship for obtaining blister copper. The final product is sent to the clients.
 
Orebody is stockwork type. Table 1 represents lithofacies series for the working area. Physical and mechanical properties of the formation were measured in the laboratories of ITU Mining Engineering Department. The tests are conducted in accordance with ISRM suggestions. Average values obtained from tests are presented in Table 2. In this study, the experimental blasts are realized in ore-dacite formation.
Stratigrafic Units - Lithology
Thickness

Upper dacite series

·Andesite-basalt, dike and silt
·Marl and limestone series
     ·Rhyolite rhyodacite
 

Upper basic series

·Andesite aglomerate                                                                
·Tuff, sandstone, limestone,
 dacitic pumice tuff
      ·Volcanic conglomerate
500 m

Lower dacite series

·Dacite green                                                                                 
·Tuff, sandstone, limestone
 (ore zone / mineralization)
     ·Tuff (ore zone / mineralization)
800 m
 
Table 1: Lithofacies series for the working area
 (Eskikaya et al., 1991)
 
Formation
Uniaxial Compressive Strength (kg/cm2)
Uniaxial Tensile Strength (kg/cm2)
Modulus of Elasticity (kg/cm2)
Density
(gr/cm3)
Altered Purple Dacite
337
40
47,500
2.58
Purple Dacite
621
75
86,300
2.64
Green Dacite
477
66
37,420
2.60
Ore-
Dacite
655
79
52,300
2.85
 
Table 2: Mechanical and physical properties of the rocks in the working area (Eskikaya et al., 1991)
 
DRILLING
 
In Murgul Copper mine, a DTH (down the hole) drilling machine is used with 165 mm drill bits. Drill bits are button type. Hole length is 14 m. The drilling time of each hole is between 90-115 minutes. Bit life is 210 m. Because of the abrasive structure of the rock formation bit life is quite short. After one hole is drilled, the bits are sharpened. In site diameter of all the bits are measured and sharpening numbers are recorded. Table 3 represents actual diameter of the 165 mm drill bits. As shown in table, each sharpening cause average 1,96 mm diameter loss. Actual diameters of the bits are average 151,7 mm.
 
This beat wear also affects the blasting performance. The blasting pattern is determined according to 165 mm hole diameter. But, because of the beat wear all holes are drilled under 165 mm diameters. This means lack of explosive during blasting in comparison to calculated blasting pattern.
 
Figure 1 shows the relation between hole diameter and total explosive amount used for 30 holes blast in Murgul Copper mine. Total explosive amount used for each blast is drawn for different beat wear situations. Calculated blasting pattern for 165 mm hole indicates 7692 kg explosive usage. Every 1 mm diameter loss causes nearly 93 kg less explosive usage. For example, if the holes drilled 152 mm in diameter, it causes lack of explosive nearly 1164 kg. No doubt, this situation affects blasting performance negatively.
Bit No
Sharpening number
Actual diameter (mm)
Diameter Loss for
each sharpening (mm)
314
2
162
1.5
305
8
149
2
308
8
148
2.12
309
7
149
2.28
328
13
142
1.77
310
7
150
2.14
364
9
149
1.78
374
3
158
2.33
357
2
160
2.5
355
11
144
1.91
368
7
153
1.71
358
6
156
1.5
Average
 
151.7
1.96
 
Table 3: Actual diameter of the drill bits
 
 
Figure 1: The relation between hole diameter and total explosive amount used for 30 holes blast
 
BLASTING ACTIVITIES IN MURGUL
 
In blasting operations, emulsion explosive (Emulite 1200Ô) is used as column charge. Priming is also emulsion explosive (Emulite 100Ô). Emulsion explosive is chosen due to its high efficiency in Murgul formation. Excellent water resistance of emulsion explosive is another important advantage (Table 4). Non-electrical milisecond delay system (Nonel Unitedä) is used to initate the carge. The timing pattern is designed as 42 ms delay between holes and 109 ms delay between rows.
 
Inhole, two primers are used. First primer (inhole detonator 475 ms delay) is located at the bottom of the hole. The second primer (inhole detonator 500 ms delay) is located under the stemming. If the primer at the bottom fail to initate, the upper primer initates the charge. There is not any need for secondary blasting. The boulders in muckpile are broken into pieces by a hydraulic excavator. All blast design parameters are summarized in Table 5. Figure 2 shows initiation sequence a typical blast performed in Murgul.
Specifications
Emulite 1200Ô
Emülite 100Ô
Density
1.20 gr/cm3
1.15 gr/cm3
Detonation velocity
5,500-6,000 m/sec
5,500-5,800 m/sec
Energy
2.2 MJ/kg
2.8 MJ/kg
Sensitivity
Booster sensitive
Cap sensitive
Water Resistance
Excellent
Excellent
Weight-Dimensions
(bulk)
2 kg-90x270 mm
Table 4: Technical specifications of the explosives
 used in Murgul
Hole Diameter
165 mm
Hole Length
14 m
Subdrilling
2 m
Hole inclination
900
Burden
5 m
Spacing
5 m
Hole Pattern
Square
Charge Depth
10 m
Bench Height
12 m
Number of Blastholes
30-40
Number of Rows
3-4
Specific Charge
0.85 kg/m3
Specific Drilling
0.046 m/m3
Stemming
4 m
Table 5: Standard blast design used in Murgul
 

 

 
Other fragmentation conditions remaining constant the design of a blast rests principally on three parameters. They are burden, spacing and specific charge (Dojcar, 1991). In this study, appropriate blasting design parameters were investigated using the approaches of blasting researchers. Different 8 approaches compiled by Jimeno (1995) and Arıoglu (1990) are used to determine burden and spacing values for Murgul Copper Mine. The burden and spacing values, which are determined from empirical formulas, are presented in Table 6.
 
Burden (m)
Spacing (m)
Bhandari
4.81
5.24
Atlas Powder
4.13-5.78
4.13-7.43
Konya
4.64
5.56
Olofsson
5.7
7.12
Arioglu
4.7
4.81
Naapuri
4.12-6.6
5.15–8.25
Hagan
3.3-4.95
6.6
Ash
5.78
5.78 -11.56
Table 6: Calculated burden and spacing values for Murgul Copper Mine.
 
The smallest burden values (4.64 m and 4.7 m) are obtained from Arioglu’s (1990) and Konya’s (1990) equation. The largest value (5.78 m) is obtained from Ash’s (Bise, 1986) equation. Atlas Powder (1987), Naapuri (1988) and Hagan (Ozer, 1996) suggest different ranges from 3.3 m to 6.6 m. For blasthole spacing the smallest value (4.81 m) is obtained from Arioglu’s equation and the largest value is obtained from Olofsson’s (1990) equation (7.12 m). Ash suggests the widest range from 5.78 m to 11.56 m.
 
Four drilling patterns are chosen based on the results of Table 6 and the field observations. These are 4.5x4.5 m, 4.5x4.75 m, 5x5 m (currently applied pattern in Murgul) and 5x5.5 m. The other design parameters are kept constant through the tests: subdrilling of 2 m, stemming of 4 m and hole inclination of 90°. Then, experimental blasts are performed. Table 7 shows blast design parameters for field-testing.
Drilling pattern (m)
(4.5x4.5)
(4.5x4.75)
(5x5)
(5x5.5)
Specific charge (kg/m3)
1.06
1.00
0.85
0.78
Specific drilling (m/m3)
0.058
0.055
0.046
0.042
Subdrilling (m)
2
2
2
2
Stemming (m)
4
4
4
4
Column charge (kg)
256.4
256.4
256.4
256.4
 
Table 7:  Blast design parameters for field-testing
 
 
 
 
 
 
 
 
 
 
 
 
 
 
 
 
 
 
 
 
After each blast, a visual analysis is performed to evaluate blasting efficiency. Displacement of muckpile, presence of misfires is observed. All boulders are counted and their dimensions are measured. There is not any excessive backbreak problem for all four patterns. Absence of the backbreak is one of the indicators of the adequate inter-row delay time. It is known that there is an almost complete absence of backbreak where adequate inter-row delay times are used (Oliver H. Phililp, 2003).
 
PARTICLE SIZE DISTRIBUTION
 
Leica Qwin Image analysis programme is used to determine particle size distribution. At least four pictures are taken for each shoot to represent muckpile properly. Additionally, the boulders (fragments more than 1 m. in diameter) are counted. Table 8 shows particle size distribution and number of the boulders larger than 1 metre. Figure 3 represent a sample image used to determine particle size distribution.
 
Pattern (m)
d50
Particles Between
3-20 cm
Particles >40 cm
Particles >1 m
4.5x4.5
7.28
81.5
3.05
32
4.5x4.75
7.36
70.87
5.51
56
5x5
7.55
82.6
3.66
65
5x5.5
7.71
72.3
5.64
71
Table 8:  Particle Size distribution
Figure 3:  Sample image used to determine particle size distribution
SHOVEL LOADING RATE
 
In second step, shovel loading rate is measured in order to determine muckpile loading efficiency. Loading time is recorded during loading operation for each experimental blast.
This method of fragmentation assessment assumes that the faster the mucking the better the production (Singh B.&Pal Roy P. 1993). The muckpile of the first shoot (4.5 m x 4.5 m) is determined as the fastest loaded muckpile. In Table 9, first column shows loading cycle duration, second column shows rock digging duration, and third column shows shovel loading capacity hourly. Loading Capacity (A) is calculated the formula given below.
A =  (1. Saltoglu S. 1992)
v = Volume of the bucket (m3)
n = Bucket filling factor (0,90)
i = Working yield (50/60)
k = Factor of rock looseness
p = Duration of cycle (sec)
 
Blasting Pattern
(m)
Cycle durations (sec)
Digging duration
(sec)
 Loading capacity (m3/hr)
(4.5x4.5)
41.56
14.2
190.1
(4.5x4.75)
42.5
14.6
185.9
(5x5)
43.75
17.4
180.6
(5x5.5)
44.2
15.3
178.8
Table 9: Shovel loading rate measurements
 
COST ANALYSIS
 
Detailed cost analysis is performed to determine unit drilling-blasting costs. Figure 4 shows unit drilling-blasting cost of the different drilling patterns As seen in Figure 4, widening of the drilling patterns provides the reduction of unit cost of drilling – blasting.
Figure 4: The unit cost of the drilling-blasting
 
ENVIRONMENTAL EFFECTS
 
OSMRE (Office of Surface Mining and Reclamation) Regulations may be used to determine flyrock distance. According to the regulation; flyrock travelling in the air or along the ground shall not be cast from the blasting site, more than one-half the distance to the nearest dwelling or other occupied structure (30 CFR Sec. 816.67 c). There should not be a flyrock stuation beyond the controlled by authorised personnel. There  are  different  approaches  used  to determine maximum rock throw length in the literature. In this paper, the equation developed by SVEDEFO (Swedish Detonic Research Foundation) is used to predict maximum rock throw length for Murgul copper mine. The equation makes it possible determining maximum throw length and the size of the fragments travelling in the air.
Lm = 260 x d2/3     (2. Jimeno C.L et al, 1995)
Lm = Maximum throw length (m)
Lm = 260 x 6.52/3
Lm = 906 m
φ = 0.1 x d2/3     (3)
φ = Size of rock Fragments (m)
d = Hole diameter (inch)
φ = 0.1 x 6.52/3 = 0.35 m
 
The residential areas near mine are closer than calculated distance. Nevertheless, during site investigation, there is no dangerous flyrock situation.
 
Instantel MiniMate Plus model vibration istruments is used in order to measure vibration and airblast levels. The instrument consist of three geophones (transversal, vertical, and longitudinal), a microphone, memory unit and battery. The instrument can be used for a single shot or continuous mode. Instantel MiniMate Plus records peak values of particle velocity up to 254 mm/s in three directions and it records airpressure between 88-148 decibels.
 
The instrument calculates the peak particle velocity, peak acceleration, peak displacement, and zero crossing frequency for each of three axes. After recording operation, instruments own data analyses software is used to access and analysis recorded data.
 
Table 10 (at the bottom of the text) indicates peak particle velocities for recorded blast. Table 11 shows Maximum permitted particle velocities at various distances from blast site (OSMRE 30CFR Section 816.67(d)(2)(i))
 
As seen in Table 10, particle velocities are quite low. There is no risk for structures near the mine area. The main factors, which are effective on occurrence of the low particle velocity, are usage of the NONEL initiation systems and low charge amount per delay.
 
Distance
[m]
Maximum particle velocity
[mm/sec.]
0-91.44
31.75
91.74-1524
25.4
>1524.30
19.05
Table 11: Maximum permitted particle velocities at various distances from blast site (OSMRE)
 
DISCUSSION AND CONCLUSIONS
 
The greatest tonnage was produced with the drilling pattern of 5x5.5 m. It is possible to reduce explosive consumption up to 8-9%, and amount of drilling up to 11-12% with the 5x5.5 m drilling pattern relative to 5x5 m current mine pattern. But this pattern produces many boulders. Image analysis put forward that reduction of the burden and spacing values provides finer fragment size and decreased boulder amount in muckpile. 4.5x4.5 m pattern provides the best blast performance. This pattern provides reduction of 50% in amount of the boulders in comparison with currently applied pattern in Murgul. Also the muckpile of 4.5x4.5 m pattern is determined as the fastest loaded muckpile. Application of the 4.5x4.5 m blasting pattern would cause an increase in unit cost of drilling- blasting by 12%. However, the excavator wastes about 20% of its working time to separate the boulders to prepare a working area for itself; the implementation of the 4.5x 4.5 m pattern in place of the current mine pattern of the 5x5 m would provides faster haulage operation. Also, there would be a reduction of 45%-50% in secondary breakage costs. Application of the 4.5x4.5 m blasting pattern instead of current mine pattern of the 5x5 m would provide more efficient and profitable production.

 
 
Pattern (m)
Distance  (m)
Total Charge (kg)
Charge per delay (kg)
Maximum Particle Velocity (mm/sec) / Frequency (Hz)
Transversal
Vertical
Longitudes
Peak Vector Summary
4.5x4.5
350
6436
256.4
2.03/10
1.65/10
2.67/8.8
3.22
4.5x4.75
135
10640
512.8
19.4/34
17.9/37
22.5/43
28.3
5x5
255
8820
512.8
1.52/17
1.90/43
1.90/37
2.35
5x5.5
210
8256
512.8
11.2/
12.7/
11.3/
14.8
Table 10: Peak Particle velocities for recorded blasts
 
Figure 2: Initiation sequence of a typical blast performed in Murgul.
 
 

 
REFERENCES
 
1.        Eskikaya, S. et al, 1991. Geostatistical Investigations of KBI Cakmakkaya and Damar Open-Pit Mines for Slop Stability, Research Project, İTÜ Research Center of Earth Since
 
2.        Dojcar, O., 1991. Investigation of Blasting Parameters to Optimize Fragmentation, Trans. Instu. Min. Metall. (Sec. A: Mining Industry), Jan-Apr 1991. p. 100
 
3.        Jimeno, C.L., Jimeno, E.L., Carcedo, F.J.A., 1995. Drilling And Blasting of Rocks, A.A.Balkema, Rotterdam. p. 290-291, 366-367
 
4.        Arioglu, E., 1990. A Semi Analytical Approach In Order To Determine Burden Values in Surface Blasting, II. National Rock Mechanics Symposium 5-7 Nov, Ankara. p. 55-83
 
5.        Bise, Christopher j., 1986. Mining Engineering Analysis, Society of Mining Engineers, Colarado, USA. p. 140
 
6.        Konya, C.J., Walter, E.J., 1990. Surface Blast Design, Prentice Hall, New Jersey. p. 100-101, 266
7.        Atlas Powder Company, 1987. Explosives And Rock Blasting, USA. p. 245, 297-300, 347, 367.
8.        Naapuri, J., 1988. Tamrock - Surface Drilling and Blasting, USA. s. 285, 359,371
 
9.        Ozer, U., Anil, M., 1996. The Empirical Investigation Of The Approaches Used For Drilling-Blasting Design, 2. Drilling and Blasting Symposium, .16-18 Jan, Ankara. p. 107-113
 
10.     Olofsson S.O., 1990. Applied Explosives Technology For Construction And Mining, Applex, Sweden. p. 22-24, 63-73, 110-112
 
11.     Bhandari, S., 1997. Engineering Rock Blasting Operations, A.A.Balkema, Rotterdam. p. 22-23, 94-95, 193-199
 
12.     Oliver, Philip H., 2003. Changes to drill pattern and adequate inter-row delay time improve blasting performance, CIM Bulletin, Vol 96, No 1071, p 60-65
 
13.     Singh, B. & Pal Roy, P., 1993. Blasting in Ground


Excavations and Mines, Balkema, Roterdam. p. 114-116
 
14.     Saltoglu, S., 1992. Open-pit Mining, Istanbul Technical University Press, Istanbul, p.117
 
15.     OSMRE, Code Of Federal Regulations, 30 CFR Sec. 816.61 - 816.68, 2005. http://www.osmre.gov/rules/subchapterk.htm#31
 

tags: QAZVIN