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By Nural Kuyucak, Graham A. Johnson

Noranda's Geco Division is presently developing a comprehensive reclamation plan for closure of its Manitouwadge, Ontario operation, expected late in 1995. As part of a program of progressive rehabilation, the existing 20 year old water treatment plant will be upgraded from a low density to a high density lime neutralization process treating AMD and excess site water. The process, developed by the Noranda Technology Centre, utilizes a two step rather than a single step process to bring the Ph to the required neutralization level, a variation which contributes to improved sludge properties such as higher density and decreased viscosity. This is of particular significance where sludge pumping or storage volumes may present a concerns.

Jan 1, 1995

By Tracy-Lynn Humby

A recent report commissioned by the South African Chamber of Mines, Mining Vision 2030, identifies the South African mining industry?s negative legacy, its exploitation of people and the environment, as the single biggest hurdle to maximizing the industry?s impact on national development. Industry leaders have identified the need to put the past ?to rest?. With reference to three recent cases involving members of the South African mining industry, this paper points to the continuities between the past and present. It argues that the metaphor of ?multiple crossings? is more generative than the notion of putting the past ?to rest? for purposes of engaging with mining?s negative legacy.

Aug 1, 2013

By E. Laatio, W. McKay

"History of Tara Mines Ltd.Tara Exploration and Development Company Ltd. was formed in Canada by Irishmen in 1953. The discovery of an orebody approximately six miles from the historically famous Hill of Tara in County Meath was a wild dream . Nevertheless this dream became a reality. Tara Mines Ltd. was formed, and the mine developed some 35 miles northwest of the City of Dublin in the Irish Republic.Tara's original drilling program indicated a zinc-lead orebody of approximately 60 million tonnes grading at 10.1% Zn and 2.6% Pb. The terms of the mining lease stipulated that the Irish government would acquire 25% equity interest in the operation. The mine was thus owned 75% by Tara Exploration and Development Company Ltd. of which the principal shareholders were Noranda and Northgate, together with 25% being held by the Irish government. The mine was purchased in March 1986 by Outokumpu Oy, a Finnish mining and metallurgical company, and together with the 25% government stake are the current owners. Outokumpu Analysis on Tara Mines Ltd.Upon ownership Outokumpu carried out a comprehensive analysis on the operation and came to the following conclusions :1. Tara had a good orebody;2. Tara had a skilled, capable, stable workforce;3. Tara's concentrator and processing equipment was adequate to meet requirements;4. Tara's mobile equipment fleet was too old and too numerousin numbers; and5. Tara had too many employees, both staff and hourly paid .With ore reserves of approximately 50 million tonnes and at a planned extraction rate of 2.6 million tonnes it meant that the mine had a lifespan of approximately 20 years. To be a viable operation into the future, the mine had to become more efficient. Operating costs and manpower levels had to be dramatically reduced."

Jan 1, 1989

By K. A. Foxford, D. A. Polya, R. P. B. Hebblethwaite

"Abstract-A variety of techniques for interpreting systems of mineralized echelon veins is discussed using examples from Panasqueira and SW England. As aids for exploration and mine planning, and as a tool for ore reserve evaluation, particular emphasis is placed upon establishing vein morphology in the third dimension, and upon understanding vein dilation and propagation histories. Special attention is also placed upon investigating vein thickness and spacing relationships, and upon understanding vein segmentation. The relevant data for these techniques are all readily obtainable from stope and core sources.The vein systems described were produced in host-rock sequences that underwent incremental dilation in response to fluid overpressures, and individual veins display evidence for mineral infilling synchronous with their propagation. Echelon vein geometries resemble those reported from igneous dikes and sills: vein walls are stepped and veins are divided into segments arranged in echelons. Changes in vein geometry, from arrays to composite structures, reflect both the evolution of the vein system with time, and the changes occurring between the margins (arrays) and core (composite structures) of the system. These characteristic morphologies can be used to infer vein propagation and fluid injection trends, proximity of neighboring veins, and the shape of the system as a whole. The techniques are applicable to systems of echelon veins produced under tensile conditions in rocks where faults were absent at the time of vein formation, and where subsequent deformation has been minimal. Further work is required to assess the application to echelon vein systems formed within more complex tectonic regimes."

Jan 1, 1995

By F. -Z. Ji, G. A. Irons, M. Barati, K. Coley

The use of oxygen in electric arc furnace steelmaking is tied to the use of fossil fuels, and slag foaming. The current work examines the kinetics of carbon injection into EAF slags especially the saturation injection rate and available reaction area in the slag foam. Based on individual rate data and experimental gasification rates of carbon in the slags, the available reaction surface area in the gas bubbles was found to be around 1900¬4700 cm2.kg-slg-1 for carbon injection into EAF like slags at appropriate injection rates and steel making temperatures. The ratio of number of bubbles to number of particles is 1 to 17 and the weights of carbon per unit reaction area are 0.001-0.008 gC-cm-2. Using the relationship between active reaction area and x supplied in this study, the gasification rates of carbon in slags can be estimated with gas-slag interfacial reaction rates.

Jan 1, 2004

By D. Seke, E. Katsande, R. Khumalo, K. Seerane

Palabora Mining Company (PMC) processes a multi-component ore body using a circuit consisting of two AG mills in parallel, closed with vibrating double deck screens and hydrocyclones. The bottom screen undersize feeds the hydrocyclones, but the screen oversize products can be treated in a variety of ways. The top deck screen pebbles can be crushed and then combined with the bottom deck oversize and recycled to the AG mill. Alternatively both the top and bottom deck screen oversize can by-pass the recycle crusher and be recycled directly to the mill. A full tap-off is a configuration when both the top and bottom deck screen oversize streams are stockpiled. The other configuration is when the bottom deck screen oversize is recycled to the mill while the top deck screen oversize is stockpiled - termed partial tap-off. This article reports the findings from the surveys performed to evaluate the performance of the PMC auto mill circuit when operated under the four different configurations discussed. Ore characterisation tests were performed to assess the deportment of the key ore components for the different configuration.

Jan 1, 2012

By B. Farabi Khoshmehr

In the present study the effect of Silicate Sodium oxidation agent concentration and Plasma Electrolytic Oxidation (PEO) treatment time were investigated on properties of oxide layer on Aluminum alloy. The phase composition and surface morphology of PEO films were evaluated through x-ray diffraction (XRD) and scanning electron microscopy (SEM). Pin on Plate abrasive machine were used to investigate of were resistance of coating. The results showed that oxide layer which formed on Aluminum surface consist of high temperature a- Al?O? and ?- Al?O? phases. The diffraction peak intensity and percentage of these phases were increased by increasing silicate sodium concentration. Microscopic studies revealed a porous surface for PEO coating. This was affected by silicate sodium concentration and oxidation time. The wear test results showed that the weight loss of PEO coating was reduced due to improvement of their wear resistance.

Oct 1, 2011

By Mike Sapiro

"Heath Steele Mines is a base metal producer situated near Newcastle, N.B. presently treating 3500 TPD of complex Cu-Pb-Zn ore. The ore consists of massive fine grained sulfides averaging I. 1% Cu, 1.6% Pb and 4.5% Zn.The original mill started in 1956 at 750 TPD and was successively expanded through 1500 to 3000 TPD in 1969. In 1972 an expansion to 4000 TPD was initiated. It was planned to obtain the maximum benefit from equipment and manpower by adding some process equipment, instrumentation, an on stream analyser, and a process control system. The benefits from this approach could be greater than normally expected, because of the complex metallurgy of this ore.IntroductionThe crushing plant is of conventional design with two 5 ½ ft, Symons cone crushers with the crushed ore screened to -7/8 inch on a 61 x 16 1 Al I is Chalmers vibrating screen and stored in four 1000 Ion bins. There are two para I lei grinding circuits, each one consisting of the following: one Rod Mill in open circuit, one Ball Mill in closed circuit with two 20"" Krebs cyclones, and one Ball Mill in closed circuit with three 15"" Krebs cyclones, Al I the Mills are 9' x 121 Al I is Chalmers Mills. The secondary cyclone overflow from each grinding circuit is thickened and combined to form flotation feed. The final product is 65% - 400 Mesh. Two regrind mills have been installed as part of the expansion."

Jan 1, 1977

By Song Mao, Local Joint Engineering Lab of Efficient Util of Dominant Mineral Resources from Karst Regions, Guizhou Key Lab of Comprehensive Utilization of Non-metallic Mineral Resource, Xianhai Li, Qin Zhang, unjian Ye, Guizhou University, Wei Cheng, Zhihong Liu

"Phosphate tailing is a kind of industrial waste that emitted from separation processing of phosphate ores, which is usually discarged to the tailing pond without treatment. Recovery of phosphorus from tailings is of great significance for comprehensive utilization of phosphorus resource. Chemical composition of a micro-fine cryptocrystalline phosphate tailing was investigated by X Ray Fluorescence (XRF) and Inductively Coupled Plasma-Atomic Emission Spectrometry (ICP-AES), and mineral composition and occurrence of harmful elements were analyzed by optical microscope and Scanning Electron Microscope (SEM), X ray Energy Dispersive Spectrometer (X-ray EDS) and Mineral liberation Analyzer (MLA). On the basis of above, effects of those on mineral processing for phosphate tailing was analyzed. Chemical analysis results in combination with microscopy indicated that there is a certain amount of iron and aluminum and it should focus on the removal of them in separation processing. The cryptocrystalline structure, especially the micro banding structure and micro porphyritic structure indicated that there is quite amount of phosphate in fine particles and it should recovery the phosphorus as much as possible when designing the flotation technology for this difficult separation fine phosphate tailing. This study can provide theoretical guide for efficient phosphorus recovery of from fine phosphate tailing by flotation.INTRODUCTIONPhosphate rocks are vital nonrenewable resources and are essential components in agricultural fertilizers and phosphorous-based chemicals (Gharabaghi et al., 2009). About 95% of the phosphate produced around the world is consumed in the fertilizer industry (page 60 Abdel-Zaher, 2008; Jasinski, 2007). Froth flotation, a solid wettability based separation process is widely used in separation of phosphate rock and shown to be effective (Yehia et al., 1990; Singh et al., 1992; Elgillani and Abouzeid, 1993; Al-Fariss et al., 2013) .Poor selectivity and low recovery for micro-fine particles are hard issues in flotation of middle-low grade phosphate ore because Phosphate resources is becoming poor, fine and complex as the rich ores run out since the 1970s. The rapid depletion of high grade ores forces minerals industry to process low grade ores where finer grinding is required to achieve required liberation of the values, making the fine particle flotation inevitable. the performance of conventional flotation deteriorates rapidly as particle size decreases to below 10-20 µm, depending on the density of the particles (Feng and Aldrich, 1999). To date, fine particle flotation is still considered to be a challenge in mineral processing (Li et al., 2015) ."

Jan 1, 2016

By T. G. Joseph

Mines must manage variability from within their operations as well as those imposed upon them by external forces. This paper provides a framework to manage both internal and external variability through the review and adjustment of the mine plan where applicable to minimize lost production and maintain operational safety.

Nov 1, 2011

By C. Díaz, A. D. Dalvi, C. E. O'Neill

"Roast-reduction smelting (RRS) of nickel concentrate, combined with flash smelting of copper concentrate, was one of the process options that Inco considered as a means of meeting its 1994 sulphur dioxide abatement target at its Copper Cliff operations. RRS involves deep roasting the nickel concentrate in fluid bed roasters. The roaster off-gas is treated in a sulphuric acid plant, and the low sulphur calcine is reduction smelted with coke in an electric furnace to yield a sulphur-deficient matte. The latter material is converted to Bessemer matte in Peirce-Smith converters, with minimal evolution of sulphur dioxide, and the converter slag is reverted to the electric furnace. Excellent nickel and cobalt recoveries (Ire obtained. This paper discusses the development of RRS, including extensive commercial scale testing at Inco's Thompson Smelter, and its key operating parameters. IntroductionReduction of sulphur dioxide emissions has been a long-standing endeavour at Inco. At the company's Ontario Division, as shown in Figure 1, sulphur dioxide emissions were reduced from about 2 million to 685 000 tonnes per year over the period 1970-1985. Most of this impressive improvement was the result of a sustained research effort which led to increased rejection of pyrrhotite from the ore during milling while maintaining an acceptable nickel recovery. Between 1970 and 1985, the combined grade of the separate nickel and copper concentrates feeding the Copper Cliff Smelter increased from about 10070 to about 'lIJ0J0 Cu +Ni and the combined S/(CU +Ni) ratio decreased from about 2.8 to 1.65."

Jan 1, 1994

By C. B. Webber, L. S. Diaz

"The paper presents the approach taken by Craigmont Mines to grinding circuit control. A detailed description is given of a fully automatic control system which bases its control actions on the continuous measurement of particle size in the cyclone overflow stream. The control system holds a specified particle size in the cyclone overflow, and grinds rod mi 11 ore feed at increased rates. The principles of operation of the system used for continuous measurement of particle size are described, end the improvements in mill throughput and copper recovery over several months of continuous operation are presented.INTRODUCTION Craigmont Mines Limited is located in the south-western interior region of the province of British Columbia, approximately 230 road miles north-east of Vancouver, and ten miles west of the town of Merritt. A moderate climate, good roads and service facilities provided by the Canadian Pacific Railway, Inland Natural Gas Company, and the B.C. Hydro and Power Authority, have provided a very favorable mine and plant site.The preliminary metallurgical design of the Craigmont mill was based on the test results of samples obtained from the original 2500' level drift during 1959. Pilot plant runs and other research testing for design pur¬poses were conducted at the Canadian Exploration Ltd. Laboratory, at Salmo, B.C. The design flowsheet was altered in progressive stages from 2,000 to 3,000, and finally 4,000 tons per day as the ore reserves became better defined. Construction work on the 2,400' level mill site commenced in late 1960. The official opening of Craigntnt Mines Limited took place on September 15, 1961, a little over two years from the time of the metallurgical assessment of the original ore samples.Mill throughput to date has varied widely both on a daily and a monthly basis, due to the varying hardness of the mine ores. The grindability af the ore appears to be related to the oxide iron content, which varies from lows of 9% to highs of 20%. Daily tonnage has ranged from 5,000 tons per day with 10% iron in the feed, to 6,000 tons per day with 20% iron feeds."

Jan 1, 1973

By J. Desroches, J. Burghardt, B. Lecampion, A. P. Bunger

"We present detailed comparisons between theoretical predictions and experiments for the initiation of radial hydraulic fractures transverse to a wellbore. Most of these experiments provide a measure of the time evolution of pressure, length, opening and flux entering the fracture during the test. We compare all these measurements simultaneously to our theoretical predictions, thoroughly testing hydraulic fracture initiation theory, which is shown to be predictive. These comparisons notably emphasize the difference between the initiation pressure (the wellbore pressure for which the fracture initiates) and the breakdown pressure (a misnomer which simply corresponds to the maximum pressure recorded during the test). INTRODUCTION In this paper, we will compare hydraulic fracturing laboratory experiments performed in different tight materials (PMMA, cement, shale) with the predictions from a numerical solution of the governing equations of hydraulic fracture mechanics (e.g. Geerstma & de Klerk, 1969; Detournay & Peirce, 2014). We focus on transverse radial hydraulic fractures initiating from a radial notch at the wellbore wall (see Figure 1). In particular, we will clarify the difference between the initiation pressure (the wellbore pressure at which the fracture initiates) and the breakdown pressure (a misnomer which simply corresponds to the maximum pressure recorded during a test, usually carried out at constant pump rate). This difference stems from the strong viscous effects at play during wellbore pressurization, fracture initiation and early propagation (i.e. fluid entrance in the newly created fracture). It has been observed on a number of experiments (e.g. Weijers, 1995; Zhao et al., 1996) and has a strong practical application, first because the maximum pressure is the most obvious value to report and second because that maximum value is necessary to specify injection systems."

Jan 1, 2015

By G. C. McKerrow

"Mr. G.C. McKerrow, Assistant Manager of Noranda Mines and formerly Smelter Superintendent attended the meeting to answer questions as to why smelters imposed penalties for too much zinc in a copper concentrate or too much copper in a zinc concentrate, etc.D.W. Bridges -Mattagami Lake MinesWhat limits do you set on zinc and lead in a copper concentrate?G.C. McKerrow - At Noranda, we allow 5% zinc free and charge 25 cents for every unit above that. For lead, it is 1% free and 25 cents per unit above that. Lead is a much more harmful impurity than zinc because it tends to recycle more in the smelter circuit and creates a greater health hazard in the plant.E. Thornton - Sullivan GroupWHat harm does fluorine do in the smelting process?G.C. McKerrow - Fluorine usually does not occur in harmful concentrations in non-ferrous smelters. Fluorides are present in exhaust gases in the aluminum and phosphate industries. If these fumes escape into the atmosphere, they may settle out on forage grasses in concentrations that are harmful to livestock.Thorton - Sullivan GroupI have asked this question other places and have been told that it is hard on brick linings and causes gold and silver to be lost up the stack.G.C. McKerrow - Almost any volatile constituent in a smelter charge will aid in carrying other metal values. To this extent it may be harmful.E. Thornton - Sullivan GroupMr. Tait mentioned before a smelter where they have a copper-tin ore and they pay for copper, tin, bismuth and silver but object strongly to indium in the concentrate. Why would this be?G.G. McKerrow - 1 can't see why indium would be a serious impurity. It does tend to concentrate in flu dusts and if you get enough of it you can recover it and sell it but I have never heard of anyone putting a penalty on it.Question - I once ran into a small copper mine in Peru that had cadmium in the copper concentrate but no zinc. The copper smelter charged a penalty for the cadmium. Why would this be?"

Jan 1, 1971

By N. C. Clarke, T. Wu, M. le Page, K. J. Henley

In 1994, Gold Mines of Australia acquired the rights to the historic Mt. Lyell copper mine, and commenced a feasibility study into expanding production. The ore was shown in laboratory tests to be amenable to single-stage, fully autogenous grinding (FAG) of underground primary crushed ore, and this route offered the maximum cost savings. Flotation performance was expected to be enhanced. However, batch flotation tests done on pilot-scale autogenously ground ore showed an unexplainable recovery loss of between 1% and 3% Cu. A second pilot plant trial was set up, with the ground product feeding banks of continuous flotation cells. Single-stage FAG, FAG/pebble and FAG/ball circuits were tested. Results confirmed a loss of about 3% recovery with single-stage FAG milling. The FAG/ball circuit eliminated the loss. Analyses showed that autogenous milling probably produced more chalcopyrite/pyrite composites and that recovery of these composites was lower than with ball milling. In size fractions below about 53 microns, where chalcopyrite is well liberated, copper recovery was less affected by milling method and flotation selectivity. It was concluded that for the Mt. Lyell situation, a FAG/ball circuit was preferable to single-stage FAG.

Jan 1, 2002

By R. M. Oliver

THE Falconbridge property is almost completely covered with over-burden to a depth varying from 60 to over 200 feet. This overburden consists of grave], sand, and boulder, with bands of very fine silt, and there is usually a layer of water-bearing quicksand adjoining the r:ock surface. In previous shafts sunk at Falconbridge, this water-bearing quicksand had given considerable trouble, and further difficulties were expected in sinking No. 5 shaft, due to its greater size. The method described in this paper is the one adopted in sinking this shaft, after the matter had been carefully considered. No. 5 shaft is located 2,400 feet east of the old main No. l shaft, and is sunk in norite in the hanging-wall of the ore-body, which is, however, ? practically vertical to the deepest known intersection. The greatest depths of overburden prevail on the south side, or footwall, of the ore-body, and this, together.. with the fact that the footwall is wet and more badly fractured, swayed the final decision to sink in the hanging-wall. After the shaft site had been chosen and cleared, a diamond drill was set up a few feet away and a hole bored to a depth of 405 feet to test the over-burden and character of the norite bedrock. The depth of overbudren indicated by the hole was Ill feet, with the last 10 feet water-bearing. The norite was rusted and broken up for only a few feet below bedrock. A steel sinking shield (Figure l) was assembled at the shaft location, and consisted of 5/8-inch plate, 5 feet deep, built around a frame of 6 inch by 6 inch angle-iron, and braced at the corners with the same material. Twelve-inch I-beams held the shield as rigid as possible. In order to reduce binding against pebbles to a minimum, the shield was constructed with a 1-inch batter on all sides.

Jan 1, 1938

By Fred G. Hewett

The Cassiar orebody is located in a sill-like serpentinite body intrusive into the Devonian-Mississippian sedimentary rocks of northern British Columbia. Subsequent intrusion of the Cassiar batholith formed the McDame synclinorium and produced chrysotile asbestos. Present information indicates an orebody of approximately 16,277,000 tons amenable to open-pit mining, but continued diamond drilling will be needed to accurately define total tonnage and grade. A program has been developed at the mine to provide the required information on a daily basis, and work is continuing on more effective dollar-value prediction methods.

Jan 1, 1978

By K. Matsuura, M. Ohno, T. Tokunaga

"Recently, weight saving of automobiles has been strongly required to solve the global environmental problems. Mg alloys are one of the promising candidates as a new material for automobile components. However, the application of the Mg alloys is limited due to their poor corrosion resistance. In the authors’ preliminary study, an Al/Mg/Al clad sheet was fabricated by hot extrusion and subsequent hot forging. The clad sheet exhibited a high corrosion resistance and excellent super plasticity. More specifically, the elongation of the clad sheet was as high as 550%.Importantly, despite that the elongation of Al itself was only 40%without cladding with Mg, the ductility of Al was greatly improved in the form of the clad sheet. In the present study, in order to clarify the mechanism of the superplastic deformation of the Al part in the clad sheet, the microstructure evolution of the Al during the tensile de formation is investigated. The tensile tests were conducted at a temperature of 573 K and at a strain rateof1.0×10-3s-1. The Al exhibited fiber-type texture at the beginning of the tensile deformation, and brass-type texture after the elongation of 50%.INTRODUCTIONMg alloys have been gaining attention because of their excellent properties e.g. light weight, good recyclability and high electromagnetic shielding. However, their practical applications have been limited due to their poor corrosion resistance. In the authors’ previous study, in order to solve this problem, Al coating on Mg alloy by hot extrusion has been proposed, and the Al-coated Mg alloy plate was fabricated(Tokunaga, Matsuura, &Ohno,2012). The plate was hot-rolled and an Al/Mg/Al clad sheet was fabricated. No cracks or debondings were observed at the Al/Mg interface. The sheet exhibited a high corrosion resistance in a comparable level as pure Al. Also, the sheet exhibited a large and uniform elongation of 550%in a tensile test at a temperature of 573 K and at a strain rate of 1.0×10-3s-1without any breakings or debondings of the Al part(Tokunaga, Matsuura, &Ohno,2014).Mg alloys have been known to exhibit a superplastic characteristic when the grain size is small. However, it is known that the pure Al does not generally exhibit the super plasticity. As mentioned above, the Al-coated Mg alloy sheet exhibited 550% of elongation in the authors’ previous study (Tokunagaet al., 2014). However, the pure Al itself showed only 40% of elongation, while the Mg alloy itself exhibited 800% of elongation. Thus, it has been demonstrated that the pure Al was super plastically elongated when it was bonded with the Mg alloy. Generally, deformation mechanism of the super plasticity of the Mg alloy is known as a grain boundary sliding. On the other hand, the mechanism of the superplastic elongation of the pure Al is still unclear. The purpose of the present study is investigation of the deformation behavior of the Al part in the Al/Mg/Al clad sheet during tensile de formation."

Jan 1, 2018

By Jae-Chun Lee, Sadia Ilyas

We investigate the technical feasibility of bioleaching of metals from low grade sulphidic ore. Experiments were conducted with an adapted consortium of Sulfobacillus thermosulfidooxidans and Thermoplasm acidophilum in batch bioreactors (1.5 L), small columns (10 kg ore load, 8 % S0 sludge) and large columns (80 kg ore load, 12 % S0 sludge). Approximately 83 % Ni, 85 % Cu, 55 % Fe and 81 % Zn was leached out in batch mode while 93 % Ni, 89 % Zn, 81 % Cu and 65 % Fe was leached out in column reactors. Bioleaching potential improved by acid pre-leaching, adaptation of microbial consortium and periodic bleeding of effluents.

Jan 1, 2014

By Muhammad Zaka Emad

Cemented rockfill (CRF) is a mixture of rock aggregate and binder slurry. It is a stiff backfill, it is fast curing, its operation is simple, placement rate is reasonably high and no drainage is required. The main demerit of CRF is failure of exposed faces causing ore dilution, which can be associated with the fill placement method. This paper discusses the effects of two placement methods being practiced by mines and numerical modelling is used as a tool to evaluate stability of the backfill column. In the first method, practiced in longitudinal retreat mining of a narrow vein deposit, the rockfill is hauled and dumped into the stope with a LHD scooptram. Once dumped, cement slurry is sprayed into the stope. In the second method, a raise situated centrally above the mined stope is used to convey already mixed CRF into the stope. Stopes are mined out transversely in primary and secondary blocks in a sublevel stoping system. Problems of poor mixing in combination with both placement methods are discussed. The two backfill placement methods are analyzed for blast induced vibration damage using the dynamic Flac3D modelling technique accounting for rock mass and fill damping characteristics, and a pressure wave is used to simulate the blast vibrations. Fill heterogeneity has also been considered for the two cases and results are compared.

Jun 1, 2012