فهرست مطالب moein bahadori

Ground vibration is one of the detrimental effects associated with blasting that can damage the surrounding environment and nearby structures. In the Gol-Gohar mine in Sirjan, due to the surface expansion, the distance between the structures and the blasting blocks has decreased, leading to vibrations reaching the processing plant complex. These vibrations, by triggering sensors installed on the mills, cause power outages in the circuit, thereby increasing production costs. One solution to mitigate the waves reaching the processing plant complex is to excavate trenches along the wave path. These trenches, by creating conditions like a free surface and reflecting the waves, reduce the transferred wave energy and can prevent unnecessary shutdowns of the concentration circuit due to increased vibration amplitudes. In this study, using the discrete element software UDEC, the results of a field blasting operation were first validated, and based on the validated model, the impact of trench excavation on the propagation of blast waves was analyzed. Ultimately, the optimal dimensions of the trench, which maximizes energy absorption, were determined. According to the numerical analysis results, the excavated trench on each side of the structure should be more than 2m longer and excavated at distances greater than 3m from the structure. Meanwhile, the thickness (width of the trench) had no significant effect on wave attenuation. This trench can reflect approximately 60% of the blast waves.

Summary In this paper, the capability of foundation and support for the Sardasht dam is investigated. Finally, according to the results of numerical modeling in UDEC software, the diameter and distance of grout injection boreholes and dimensions and angle of seal curtain of the dam foundation dam have been suggested.   Introduction The injection is a process whereby a cement slurry is pushed into rock formations through a borehole, thereby reducing the permeability and deformation of the rock mass and increasing its resistance. The rocks are almost impermeable and rock mass permeability is often a function of discontinuity systems. Due to the different and unpredictable behavior of rock masses, there is no specific law for determining the distance of injection holes and generally relies on the experience and judgment of the design engineer when deciding on the borehole distance. So, it is proposed to use numerical methods in predicting the radii of injection for grouting material and also determining the ideal spacing between adjacent holes. As the main results of this study, the optimum pattern for drilling sealing systems for different locations of the Sardasht dam was determined and compared to the empirical models using the discrete element method in UDEC. The optimum deviation angle of the holes was investigated, too.   Methodology and Approaches The Sardasht dam is a trench with a clay core with a height of about 106 meters and a length of 280 meters. The total embankment volume of the dam body is estimated to be about 3 million cubic meters and the volume of clay core is about 516000 cubic meters. The water diversion system consists of two tunnel strings with an inner diameter of 7m and lengths of 627m and 682m in the right support and its height is about 46m. The right tunnel is used as the lower evacuator during the operation period. In order to investigate the in-situ condition of rock mass in Sardasht Dam area, rock mechanical parameters including rock mass quality index (RQD), specific gravity, uniaxial compressive strength, and geometry and resistive properties of discontinuities have been determined and measured. Then, the quality of grout injection in walls and foundation of the Sardasht dam was modeled, using numerical modeling in UDEC discrete element software.   Results and Conclusions The results of this study show that the appropriate borehole spacing for the walls and the foundation should be taken as 3m and 5m, respectively. Also, the results obtained from the numerical modeling of the optimum injection pressure in the construction area of the Sardasht dam were determined for different depths. Based on the numerical modeling results in order to minimize water leakage from the Sardasht dam foundation, the optimum angle of curtain installation should be 17 degrees.

Rock fragmentation is one of the desired results of rock blasting. So, controlling and predicting it, has direct effects on operational costs of mining. There are different ways that could be used to predict the size distribution of fragmented rocks. Mathematical relations have been widely used in these predictions. From among three proposed mathematical relations, one was selected in this study to estimate the size distribution curve of blasting. The accuracy of its estimates was compared to that of the RR (Rosin-Rammler), SveDeFo (The Swedish Detonic Research Foundation), TCM (Two Component Model), CZM (Crushed Zone Model), and KCO (Kuznetsov – Cunningham - Ouchterlony) relations. The comparison included assessing the accuracy (Regression, R) and precision (Mean Square Error, MSE) of the best possible fit between the mathematical relations to estimate the cumulative distribution of fragmented rocks that result from rock blasting in open pit mines (Miduk Copper Mine, Sirjan Gol-e-Gohar, and Chadormalu Iron Mines) using image analysis technique. The results showed that the power hyperbolic tangent function can estimate size distribution of hard rock fragmentation with more uniformity in fine and coarse-grained sizes (unlike soft and altered rocks with the non-uniform distribution in these regions), more accurately and with higher precision. Also, unlike the KCO, the absence of a second turning point for the largest block dimensions (Xm) in the proposed function, can guarantee the accuracy of estimations related to any range of inputs. Finally, due to the ability of the proposed relation to accurately estimate rock fragmentation distribution caused by blasting, the uniformity coefficient required for the relation is provided by a linear combination of the geometric blasting parameters, where R=0.855 and MSE=0.0037.

Ground vibration is one of the undesirable results of blasting operations. Different methods have been proposed to predict and control ground vibration that is caused by blasting. These methods can be classified as laboratory studies, fieldwork and numerical modeling. Among these methods, numerical modeling is the one which saves time and cuts costs since it takes into account the basic principles of mechanics and provides step by step time-domain solutions. In order to use numerical analysis in predicting the results of blasting operations, the accuracy of the output must be verified through field test. In this study, ground vibration caused by blasting in a field operation in Miduk Copper Mine was recorded using 3-components seismometers of the Vibracord seismograph and analyzed by Vibration-Meter software. Propagation of the waves caused by blasting in the mine slope was modeled using discrete element logic in the UDEC numerical software and compared to the results of the field test. Having tested the accuracy of the results obtained, the effect of primer location and the direction of detonation propagation in the blast hole on the rate of ground vibration caused by blasting was investigated. The results show that by changing primer location from the bottom of the hole to its top, the rate of ground vibration caused by blasting increases.

Controlled blasting is commonly employed in civil and mining engineering developments. This reduces maintenance and supporting system costs and improves the bench appearance. Safe maintenance of the wall and avoidance of damage caused by blasting is therefore important in all subsequent underground and surface excavations. Different types of controlled blasting nowadays include trim blasting, line drilling, cushion blasting, pre-split blasting, fracture controlled method and linear shaped charge. Although these approaches incur additional operational costs, but considering benefits associated with the safety condition necessary for such operations, increased production and faster progress resulting from more stable walls, improved waste/ore ratio, controlled ore concentration and the required size reduction of the rock for haulage and loading, these additional costs are justifiable. Controlled blasting is also common in open-pit mining, quarry mining, trenching and shaft drilling. In case spacing and charge quantity are evaluated based on engineering design principals, a uniform fracture with narrow width would result. This would dampen the transfer of the explosion wave outside the explosion block, when production blast holes are fired. In pre-split blasting, which is a more common technique often used in such operations, blast holes with smaller diameters and lesser spacing than normal production sizes, are applied in the last drilling row. The method could easily be applied to all types of rocks. However, drilling patterns and the required explosive charges should be determined based on rock mass characteristics, such as stiffness, roughness, existence of discontinuities. Also, in pre-split blasting, contrary to other methods, the controlled blast holes are fired 50ms sooner than the main production blast holes. In case this delay exceeds 50ms for whatever reason, the fracture produced by the controlled blast holes will be filled with post-explosion fragments and their ability in cushioning the explosive transmitted wave is seriously hampered. Hence, in applications of this technique for hard rock mass, it is customary to leave a proportion of the blast holes without charge. The blast holes diameter in this technique varies from 51 to 102 mm, with the diameter of charge ranging from 17 to 32 mm, hence decoupling is less than one. The explosive connection is often carried out by detonating cord and if wire and electric detonators are used, they should be fast triggered type of milli-second delay or better. Superposition of compression waves due to adjacent blasting holes lead to tension stresses, perpendicular to the direction of blast-hole lines. This results in tensional fractures within the rock mass. In this study, using UDEC distinct element software, the mechanism of crack propagation and superposition of pre-split blasting waves in three holes are investigated, and the data are compared with field data obtained on the conglomerate rock mass at Gotvand Olya dam. Blast holes of 76mm in diameter, 3m height, 8 cm spacing were drilled at the rock mass and charging included 7 Emulite cartridges with cortex blasting. Ground vibrations of 175.24 mms-1 and 77.33 mms-1 were recorded by two VIBROLOC seismometers, placed at 8m and 13m away from the blast hole centre, respectively. The results suggest that numerical simulation could be employed with sufficient accuracy for predicting presplit blasting.

Drilling and blasting have numerous applications in civil and mining engineering. However, there are many unfavorable associated side effects and hazards, such as ground vibrations, air blasts, fly rocks, back-breaks, unwanted displacements, crack formation and propagation, and extended crushed zones, all of which need to be predicted and controlled effectively. Ground vibrations caused by blasting can damage the zones in the vicinity of the explosion block and its associated civilian structures and equipment. In addition to environmental and structural damages, air blasts can irreversibly damage the health by affecting the hearing sense and mental stability of the personnel. Damages to the front face, caused by open-pit and underground explosions, not only increase the maintenance costs, but also make the appearance unacceptable. An important factor in reducing hazards in controlled blasting is the prediction of crack formation around the blast-hole and its propagation, which has been the subject of research since early 1950’s using field experiments, analytical methods and numerical simulations, paving way for many semi empirical correlations presented in the literature on this matter. As a rule of thumb, the radius of the crushed zone and the length of the radial cracks, are assumed to be in the order of 3 to 5 times and 40 to 50 times that of explosive radius, respectively. Hence, the radius of the crushed zone and the radial crack lengths were evaluated to be 10.16 cm and 114 cm, respectively. In this study, the results of the field studies from single blast-holes in the mudstones of Gotvand Olya dam were compared with several empirical correlations, using a blast-hole of 76 mm diameter, 2 m depth, 1 kg emulate 27 charge and a single instantaneous electrical cap. Two seismometers of VIBROLOC placed 8 m and 13 m away from the blast-hole recorded the ground vibration at 17.22 and 9.02 mms-1, respectively. The crushed zone radius and the radial crack length were measured to be 25 cm and 90 cm, respectively. The crack propagation and the ground vibration were compared with the field study results using a UDEC discrete element method. In the simulation exercise, the dynamic loading on the surrounding walls of the blast hole were assumed to be uniform and in radial direction. Also, the blast was assumed to take place instantaneously along the cylindrical charge and the semi-empirical relationship of Liu and Tidman was used to evaluate the maximum detonation pressure produced. The simulation results included a variation in the peak particle velocity with respect to the distance from the blast hole centre, a variation in the particle displacement, a variation in the applied stresses caused by the shock wave travelling, reflecting the stress wave from a free face. The numerical analysis indicated the crushed zone radius and the radial crack length to be 20 and 90 cm, respectively. Also, the ground vibrations at 8 m and 13 m distances away from the blast-hole were simulated to be 17.2 mms-1 and 9.27 mms-1, respectively. Amongst the empirical correlations used, Ash correlation (1963) revealed a radial crack length of 110 cm, and Essen et al. (2003) evaluated a crushed zone radius of 19 cm, indicating more accurate estimations. This study indicates that the numerical analysis used is capable of presenting acceptable accuracy.

Ground vibration is an inevitable consequence of blasting operations in open pit miningand civil engineering projects. A large amount of energy is released in every blastingexercise in the form of wave propagation, which can cause serious damage to thesurrounding environment. Safe design of blasting for underground structures andtunneling are nowadays carried out mainly based on the permissible peak particle velocity(PPV), which has become a major criterion in setting out the appropriate safety standardsenvisaged to prevent structural damage near such operations. Many published studieshave produced correlations for PPV prediction using experimental data The GotvandOlya Dam is one of the major civil engineering projects in the last thirty years in Iran, constructed as the last dam on the Karoon River with the aim of increasingenvironmentally friendly electrical energy of the country, seasonal flood control in theprovince, and a supply of agricultural water for the strategic farming grounds ofKhoozestan. The geological formations of Gotvand include the two main formations ofBakhtiyari and Aghajari. The existence of discontinuities in these formations as well asjoint sets and their low inherent strength caused a number of challenges in blasting designoperations. This dam is 30 km north west of Shoshtar in the Khoozestan province, and 12km away from Gotvand. In this study, 16 records of 3 components produced from 4 blastswere obtained using 4 seismographs of the type PG-2002 at Intac and surge tanks of theGotvand Olya Dam in order to investigate their effects on underground structures. Thethree-component seismometers were of the GS-11D type and the records were analyzedusing DADISP software. The explosion materials used were ANFO and dynamite, andthe maximum weight of the explosions per delay in the Intac and surge tank were in therange of 32.0 to 343.3 kg. The positions of the seismometers were in the range of 56.2 to 145.5 meters away from the centre of blast blocks. The permissible peak particle velocity was taken from the Korean Institute of Geology, Mining and Materials. Using the scale distance, an exponential equation based on the cubic root of the charge weight per delay and coefficient correlation of 85% was proposed for predicting the PPV. In thisrelationship, the geological coefficients of the Gotvand area were estimated to be 20.339and -3.08. Using a genetic algorithm, an improved coefficient correlation between theexperimental and predicted PPV of 89% was obtained. The results from the geneticalgorithm applied for the coefficients of a, b, and n in the equation of PPV=b(d/wn)a were -9.286849, 9.769703 and 0.745959, respectively. Considering this proposed relationship,the minimum distance of 56.2 m to the centre of the blast block and the permissible peak particle velocity of 254 mm/s for a 10 day old concrete, the maximum charge weight per delay was found to be 254.4 kg.