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By D Roughsedge, J van der Sandt, L van den Berg

Axial flow fans are commonly used as auxiliary fans in underground mines to supply air to development headings. As diesel dilution requirements increase the fans used in conjunction with ventilation bags are becoming larger in diameter to move greater quantities of air over greater development distances. It is not uncommon for mining operations to sometimes use a number of these larger auxiliary fans with high airflow capacity in parallel as primary fans for the operations. These primary fan stations are often considered temporary installations and the fan selection and fan station design is rarely optimised. These temporary installations may become permanent resulting in an inefficient or unreliable primary ventilation system. A recent study looked at the likely causes of multiple primary fan failures at one operation with axial fans installed as primary units where the fan selection and fan station aerodynamic design was far from optimal. This paper explores some of the typical mechanisms of failure such as high pressure stall, unbalanced airflow, and fan tip speed as well as fabrication material. The paper proposes a number of good practise principles that should be considered during the design of a fan station. A second topic covered is the advantages and disadvantages of using different types of fans such as axial, mixed flow and centrifugal units. Here the concept of fan specific speed and its relevance to fan selection is discussed. Finally the paper quantifies the effect of using a high pressure auxiliary fan as a primary fan in a low pressure circuit, which may increase the mineÆs operation cost.CITATION:van der Sandt, J, van den Berg, L and Roughsedge, D, 2013. Practical considerations when using auxiliary axial flow fans as a primary mine fan, in Proceedings The Australian Mine Ventilation Conference , pp 63-70 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 1, 2013

By H Hoogvliet

The Goondicum Crater industrial minerals deposit is located about 30 km east of Monto, Central Queensland, Australia. Early exploration programs in the 1990s found alluvial ilmenite along the banks of the Burnett River. This mineralisation was traced upstream to the Goondicum Crater. The Goondicum Crater is a topographic feature centred on a roughly circular 6 km diameter layered gabbro complex. Significant ilmenite, apatite, titano-magnetite and feldspar mineralisation are present in a weathered horizon up to 25 m thick. After several drilling campaigns in the Crater and subsequent feasibility studies, the project was developed by Monto Minerals Ltd, an ASX listed company, in 2007. However, the project never achieved production performance expectations and ran out of commissioning funds. The company subsequently went into voluntary administration in August 2008.In 2009, the RMM Capital group, of Brisbane, took control of Monto Resources Pty Ltd, and renamed the company Belridge Enterprises Pty Ltd (Belridge). Belridge commissioned a redevelopment study aimed at recommencing mining and processing operations. As part of the redevelopment study the authors planned a drilling campaign, managed a quality control program and developed a block model. This paper describes the drilling program, analytical procedures, quality control procedures and the preparation of the block model.The block model for the resource estimate was created using 1 m high blocks, wireframes defining the main material types, and 2.5 m thick horizontal search ellipsoids with a radial pattern. The deposit was ‘flattened’ prior to modelling. To assist with the economic evaluation of the project, each block in the model was assigned a net dollar value based on economic parameters provided by Belridge, including anticipated revenues and costs. This modelling technique allowed for efficient investigation of various scenarios and was used extensively in the development and optimisation of the project’s financial model. At a zero net value per block, comparable to a break-even cut-off grade of four per cent ilmenite equivalent, the Indicated plus Measured resource is estimated at 51 Mt, averaging 4.1 per cent ilmenite, 1.9 per cent apatite and 2.5 per cent titano-magnetite.

Oct 5, 2011

By S Bhandari

This paper is a step in the direction of developing a computer model of the near source dispersal of suspended particulate matter (SPM) in planetary boundary layer (PBL). Since Surface Mine blasting is done during daytime, study of daytime dispersion in convective boundary layer (CBL) is relevant to the mining engineers. The scope of this paper covers dispersal after the blast in any type of PBL ie stable, convective and neutral but only under weak wind condition. Mathematical analysis of the dispersion of vertical dust column has been carried out during weak wind conditions in a tropical meteorological environment. It is noted that dust concentration reaches its maximum value some distance away from the point of blast, immediately after the blast (t*=1) and thereafter rapidly decreases. After a short interval from the time of blast at t*=11 the dust concentration shows linear decrease in concentration with distance. Dust concentration also rapidly drops during the interval t*=1 to t*=11, thereafter, rate of drop is nearly the same at any distance from the point of blast.

Jan 1, 2001

"Recognising that the mining industry is wedded to the break-even cut-off grade rather than a cut-off grade that optimises value, the author examines some perils and pitfalls in the use of break-even grades that lead to what the author calls ‘negatively geared ore reserves’.Negatively geared ore reserves occur when a portion of the ore reserves actually has a negative value and is effectively being subsidised by the more profitable portion of the ore reserves. This situation is generally difficult to detect with the level of information available to the public investor. The ore reserves can still be quite profitable in total, but now possess a level of risk that may not be appreciated.The risk profile (financial risk) of a negatively geared ore reserve is far greater than for a non-negatively geared ore reserve. Like a negatively geared investment property in a falling property market (for readers familiar with the Australian tax minimisation investment strategy), when the price of a commodity falls, a company that possesses a negatively geared ore reserve loses value at a magnified rate relative to a company that has a neutral or positively geared ore reserve. In the extreme case of a major commodity price reduction, as happened with the gold price in mid-2013, the negatively geared ore reserve can easily become negative value in total, not just in part.The negatively geared ore reserve is sometimes intentional – for instance in order to increase published ore reserve tonnages and meet the CEO’s key performance indicators – but is also often unintentional, resulting from various errors in the break-even calculation. Common errors in calculating a break-even grade include using fixed metallurgical recoveries, not including allowances for ideal laboratory versus real plant results, ignoring available operational data and using modal rather than mean parameter values in cost and performance calculations.This paper examines a common value distribution profile in many ore reserves and the effect of common break-even grade errors on this profile.CITATION:Poniewierski, J, 2016. Negatively geared ore reserves – a major peril of the break-even cut-off grade, in Proceedings Project Evaluation 2016, pp 236–247 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 8, 2016

Sampling of CIL/CIP circuits is typically done by compositing samples of feed and tails over eight or 12 hour periods. As the residence time in the leaching tankage may be approximately 20 hours, the tails from any period being derived from material entering the circuit 20 hours earlier do not correspond to the head from the same sampling period. Moreover, mixing in the series of tanks in a circuit also modifies correspondence between input and output. A special purpose simulator has therefore been developed to allow for the mixing and delay in the tankage and permit estimation of the head assays corresponding to any tailings sample. Thus more accurate assessment of the day-to-day recoveries of gold and silver follows. The model, programmed in Excel, assumes that the tanks are perfect mixers. By inputting the sequential head grades and considering that no leaching has occurred, the head grade which corresponded to the tailings for each period could be estimated. Allowance is made for variations in feed rate over the sampling period by computing the average flow rate of slurry and thus the residence time of each increment of the tailings. This simple model and its application have been found to be subject to some limitations, which are discussed in the paper, resulting in some unresolved variability in results. However, for a typical circuit even the simple model was able to greatly improve consistency of recovery data with variations being approximately halved by eliminating much of the anomalous variation arising from the lack of correspondence between head and tail when changes in head grade are occurring. Therefore, this permits genuine variations in metallurgy to be identified with greater confidence and thus operational shortcomings can be identified and addressed more quickly.

Jan 1, 2006

The character of alluvial gold deposits presents severe difficulties, and sometimes limitations, on the extent to which metallurgical efficiency can be measured. Very little monitoring of metallurgical performance is practised. An indication of performance is often extracted from the comparison of production to the estimate. This is a dubious practice. The recovery efficiency of conventional practices, viz sluices, tables and jigs, falls off sharply below 0.2 mm. Field evaluation procedures tend to simulate these conventional practices. Thus the presence of fine gold may not be detected, particularly if obscured by clay. There is increasing evidence that fine gold could, in places, be present in sufficient concentrations to be economic. Most gold in paleoplacers is finer than 0.1 mm. There are established practices elsewhere in the mineral industry that may physically recover gold down to 5 microns in size and which could find application for fine alluvial gold.

Jan 1, 1984

By S Bluhm, LádeáOliveira, R Funnell

Taquari Vasouras potash mine (Vale Fertilizantes) has many interesting features including virgin rock temperatures of 50¦C at depths of 700 m. The mine has a relatively small shaft system with limited ventilation carrying capacity, which was not adequate to exploit the full extent of the deposit. This limitation was overcome by installing two large-scale recirculation systems with underground reconditioning centres. This paper describes this unique mine ventilation and cooling system and some of its history over the past two decades. The two controlled recirculation systems have been operating safely for some ten years and the use of large-scale recirculation has been a fundamental component in the increase of production and extending the mining further-and-further from the shaft centre. The recirculation systems have played a huge role in the mine profitability and comprehensively extending the mine life. The paper also discusses a number of novel design and operating features such as reverse running pumps for energy recovery. The underground chilled water system design required a number of special safety features in order to minimise the risk of water leakage since this could be catastrophic in a potash mine. This paper describes a good example of a mine in which the final ventilation and cooling system design ends up very different to what was envisaged at the greenfields stage.CITATION:Bluhm, S, Funnell, R and de Oliveira, L, 2013. Controlled recirculation and refrigeration at ValeÆs Taquari Potash Mine, in Proceedings The Australian Mine Ventilation Conference , pp 237-248 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 1, 2013

By F Clark, J T. Carswell

Ore definition for final stope planning is often a challenging prospect in underground (UG) gold mines which exhibit high geological and grade variability. Unlike the grade control (GC) scenario in an open pit (OP), the ability for the mining geologist to plan a sampling campaign which effectively characterises the underlying mineralisation is fraught with difficulty. UG sampling patterns are strongly clustered and typically not dense enough to characterise the spatial and statistical properties. Very commonly, sample volumes are not large enough to provide a reliable grade estimate of the surrounding rock. These factors critically affect our ability to accurately map out ore positions and estimate metal quantities.At Sunrise Dam, traditional UG sampling and modelling techniques have proven to be an unreliable strategy for successful stope planning and extraction. The problem was first observed during the final cut-back in the open pit for which the mineral resource model was constructed from UG diamond drilling. Open Pit grade control reconciliations were strongly positive for both tonnage and grade compared to the resource model. Underground workings missed significant quantities of ore which was laterally defined by open pit reverse circulation (RC) GC drilling. Upon examination it was realised that the cause of the problem was one of inadequate sample density and sample support, the implications of which potentially affected the entire UG mine.As part of a broader study into underground grade control procedures, detailed reconciliation of initial stope production from the GQ domain also pointed to inadequate sampling as a major constraint to effective grade control and reserve conversion.To overcome this deficiency, dense 115 mm wet RC grade control drilling and conditional simulation modelling has been successfully implemented with the intent to replicate the grade control process in the open pit. The increase in sample density and sample volume has proven to be a very effective tool for mapping out highly complex mineralisation. A marked improvement in stope definition, reserve conversion and Mine to Mill reconciliations has been demonstrated on the back of an innovative approach to UG grade control.CITATION:Clark, F and Carswell, J T, 2014. Using reverse circulation drilling to improve sampling in a complex underground gold operation – an innovative approach to underground grade control at the Sunrise Dam Mine, Western Australia, in Proceedings Ninth International Mining Geology Conference 2014 , pp 143–152 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 18, 2014

By R McDonough

DTL2 contract C916 was awarded in March 2009 for S$340 M. This contract includes two mass rapid transit (MRT) 1100?m long bored tunnels using two slurry tunnel boring machines (TBMs), with an internal diameter of 5.8?m, two cross passages and 135?m of cut-and-cover tunnels The contract also includes the design and construction of a new underground station including station entrances, subway links, architectural finishes and external landscaping. The station, located at Upper Bukit Timah Road near Jalan Jurong Kechil will also double as a civil defence shelter.The works included significant traffic management in a dense urban area in order to create the working areas whilst keeping the requisite number of traffic lanes. This involved numerous utility diversions and included the removal of a 33-year-old pedestrian overhead bridge (POB) over busy Bukit Timah Road. The station and tunnelling works were completed on time, including a ‘dock and dismantle’ TBM retrieval and subsequent breakthrough of the separating wall.This paper will detail the methods and solutions used to construct the tunnels and stations, with a focus on the requirements for construction of the station to be a civil defence shelter if required. This involved solutions to tying reinforcement and placing concrete to high standards in a restricted space, as well as the incorporation of heavy blast doors and other associated protective measures. The efforts of the management team resulted in the project being awarded first prize in the 2011 Land Transport Authority’s Annual Safety Award Convention and being named winner of the Safety and Health Award Recognition for Projects (SHARP) award at the 2011 Workplace Safety and Health (WSH) Awards. As well, McConnell Dowell won a Green and Gracious Builder Award (Excellent Category) from the Building and Construction Authority (BCA).CITATION:McDonough, R, 2014. Design and construction contract C916 Beauty World Station and Tunnels, Singapore, in Proceedings 15th Australasian Tunnelling Conference 2014 , pp 411–418 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 17, 2014

By J Euston

In the 1990s round flotation tanks were introduced as a result of the lower fabrication costs for larger cells. It was soon observed that round tanks held a number of further advantages over the traditional rectangular cells, such as better hydrodynamics, lower costs and reduced footprint per unit volume.In the last twenty years flotation cell sizes have increased almost exponentially and designs of 250 to 300m3 are a well established standard with plans for full scale pilot installations of 500m3. The option to install these large cells gives flexibility in circuit arrangements and allows efficient and economic matching of flotation lines with modern mill designs. The benefits of scale up are well known with reduced numbers of cells required, lower power costs per unit volume, etc. These benefits will be briefly reviewed with specific reference to the installation of 300m3 to 350m3 flotation cells at the Copperton Concentrator of Kennecott Copper in Utah, USA.The sophisticated techniques required to model these ‘supercells’ have led to a critical review of the fundamentals of flotation science and engineering. In addition to outlining the advantages of large flotation cells, this paper will review the science which supports, and occasionally challenges, the ongoing trend to larger flotation cells.

Aug 8, 2011

By B Murphy, T Mattsson, R Crosbie, T Maksimainen

"An Outotec e500 TankCell, with 500 m³ of actual flotation volume, was commissioned at the Kevitsa Cu-Ni-PGE concentrator in northern Finland during 2014. This flotation cell was implemented at the head of the existing copper circuit as the first Cu rougher flotation cell. This was the first flotation cell of this size to be run under production conditions by any supplier anywhere in the world.The cell has been in continuous operation since start-up in October 2014 and has provided improvements in metallurgical performance at very low specific power consumption. On average, the cell has recovered 71 per cent of copper contained in the 0.37 per cent Cu flotation feed at a concentrate grade of 17 per cent Cu. The specific power consumption for the cell (excluding blower power) is typically around 0.4 kW/m³.After start-up, the impeller speed and superficial gas velocity were varied and the impact on metallurgical performance measured. In this paper a review of the cell operation at Kevitsa is given. Furthermore, the impact of mixing and gas dispersion on metallurgical performance is presented.Citation:Crosbie, R, Mattsson, T, Murphy, B and Maksimainen, T, 2016. Metallurgical Performance of Outotec's e500 TankCell at the Kevitsa Cu-Ni-PGE Concentrator, in Proceedings 13th AusIMM Mill Operators’ Conference 2016, pp 127–134 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Oct 10, 2016

By M J. Nugus

The Geita gold mines in north-western Tanzania have produced 5 Moz of gold from an estimated endowment of 15 Moz. They are generally classified as typical Archaean orogenic banded iron formations (BIF)-hosted gold deposits in banded iron formations as described by Goldfarb, Groves and Gardoll (2001); Neall and Phillips (1987); Phillips and Powell (1993) and Neall (1987). The controls on gold mineralisation have been reassessed to conclude that Nyankanga, and we conclude that Nyankanga, the largest deposits within the group, developed during a dynamic and protracted deformation history intimately associated with several stages of hydrothermal fluid flow and alteration. The fluid-induced fractures associated with gold mineralisation are primarily manifested as submillimetre to metre-scale fractures and breccias developed preferentially in rocks that have been previously deformed and altered by silicification or seafloor metasomatism. This realisation detracts from the exploration and mining strategies that would solely concentrate on fluid pathways and associated wallrock alteration and veining solely within the BIFs as hosts for economic gold mineralisation.

Aug 22, 2011

By E J. Sellers, A Cocker

Movement of the fragmented rock mass is an inevitable consequence of open cut blasting. Uncontrolled, unaccounted movement from blasting commonly causes ore loss and dilution from waste mixing. Therefore, the prediction of the post-blast location of both ore and waste is vital in grade control and of direct interest to mining geologists. Direct measurement techniques have matured significantly over the past decade, but their use remains expensive and labour intensive. Access and safety issues are further disincentives to regular movement monitoring. What is needed is a combination of direct measurement to establish baselines and reliable predictive modelling.The W H Bryan Mining and Geology Centre of the University of Queensland has pioneered 3D blast movement modelling using ‘physics engines’ to handle in-flight collisions and predict the final resting place of blasted material. Open cut metalliferous mines are a demanding environment for this style of modelling as they can involve complex geometries and variable buffering, which leads to inconsistent confinement. A version of the model designed to deal with these complexities is under development. A well-controlled field trial was needed to validate this new version.A validation trial was completed at a gold mine in South-East Asia. The aim was to increase throughput by increasing the blast powder factor without losing grade through dilution and mixing with waste. Four blocks were blasted to identify the effect of bench height and face confinement on ore movement, fragmentation and process throughput. The powder factor and blast timing were kept constant. Movement of the broken rock was measured by placing blast movement monitors at known positions in the block and then locating them after the blast. The majority of sensors were recovered, providing a unique insight into blast movement dynamics and generating a substantial data set for model validation.The model (called ‘MetaMove ‘) was calibrated for one blast and applied to investigate the extent of grade loss and possible mitigation strategies. Conventional mining techniques that involved firing across the strike of the orebody, and no post-blast boundary adjustment of the ore/waste boundaries prior to digging, were found to cause up to 44 per cent loss of high-grade ore, depending on the grade distributions. Post-blast adjustment of the ore/waste boundaries can negate this loss. The model indicates that altering timing and ensuring that the block boundaries are well confined can further constrain the movement of ore/waste boundaries. Thus, the model can be successfully applied to minimise grade dilution and limit misclassification of ore as waste whilst using higher powder factors to increase mill throughput. Application of a calibrated model to move ore/waste boundaries prior to digging will enable the production geology team to add significant value to the mining operation.CITATION:Cocker, A and Sellers, E J, 2014. Modelling blast movement for grade control at an open cut gold mine, in Proceedings Ninth International Mining Geology Conference 2014 , pp 377–386 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 18, 2014

The Kiirunavaara apatite-magnetite orebody is about 4 km long and approximately 100 m thick sheet-like body, which lies roughly in a north-south direction and dips from the horizontal plane along its entire length, in towards the town of Kiruna. The origin of this orebody has been a matter of controversy, with both magmatic and hydrothermal origins proposed. Traditionally and for practical purposes this rich apatite-magnetite ore has been divided into two main types: low phosphorous ore (B-ore) and high phosphorous ore (D-ore). Further the low phosphorous ore is divided into the low alkali (B1) and the high alkali (B2) ore.

Jan 1, 2006

The characteristics and causes of the spoil pile failures at the Goonyella Mine are described, and research results from investigations of spoil pile instability are discussed. Some effects on spoil stability of the development of deep strip mining projects are considered and recommendations are made for relevant research.

Jan 1, 1981

By M J. Pejovic, I Miljanovic, A Milutinovic4, T Benovic, A Petrovski

The paper presents a model of a multigrade monitoring-management structure for support to decision making and management in real and extended time in mineral processing (MP). The model was created as a result of a long-lasting research, design and development of mining, computer integrated monitoring-management systems. Apart from the general explanations, the authorsÆ state in the introduction that efficiency of a system for support to decision-making and management in mineral processing is more difficult to achieve by conventional approaches, than by the fuzzy logic approach. Starting from the multiphase structure of process flows in MP, the following are process and monitoring-management oriented graphs. The graphs served as a foundation of the multigrade hierarchical fuzzy model for support to decision making and management in MP. Following the example of the monitoring-management system of the æRudnikÆ mine flotation plant, fuzzy logic implementation was shown. The key conclusions of the paper are that fuzzyfication of layers of the decision making and management model has full justification due to the complexity and dynamics of mineral processing processes, complications and often insurmountable difficulties concerning the precise description of real systems by conventional mathematical tools, variability of external influences, and the (general) non-linearity of the MP processes management criteria functions.

Sep 26, 2011

By D Willis, A Shanahan, C Jacobs, T Luxner

The application of tunnel boring machines (TBM) for reef mining is a unique practice with benefits over current mining methods given the right geology and geometry of the orebody. The Stillwater Mining Company (SMC) of Nye, Montana, USA operates two mines for obtaining platinum metals. The operation is currently using a 5.5 m diameter open-type machine for further mine development, as well as another TBM that they own.The latest 7.1 km bored tunnel, known as the Blitz Tunnel, will be routed parallel to the ore deposit in igneous mafic intrusive rock, and will be used for the life of the mine operation. It will be used to map the ore deposit, and provide a rail haulage route.Detecting the reef in relation to the TBM requires careful analysis. Diamond core drilling occurs ahead, above, and alongside the machine at every 150 m. The core is drilled, logged, and interpreted on the spot, concurrent with boring. Based on the data, the TBM is readjusted so that it stays on the correct bore path.This paper will highlight the application and benefits of TBM usage for reef mining, with detailed descriptions from the mine site.CITATION:Willis, D, Jacobs, C, Luxner, T and Shanahan, A, 2014. Tunnel boring machine usage speeds up development at Stillwater Mine, in Proceedings 15th Australasian Tunnelling Conference 2014 , pp 683–690 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 17, 2014

By M Svartsjaern, D Saiang and K Mäkitaavola

The Kiirunavaara mine is a large-scale (28 Mt/a) iron ore mine located in northern Sweden. The mine is owned and operated by Luossavaara-Kiirunavaara Aktiebolag (LKAB) using sublevel caving (SLC). The SLC area underlies an open pit decommissioned in the 1950s, the current main SLC haulage level is situated at an depth of 1100 m. The main orebody dips 60 degrees to the east with a distinct boundary defining the hanging wall to the east and footwall to the west. Since the 1980s damage has been observed both in the footwall rock mass as well as on the footwall crest of the open pit. Recent investigations indicate that the underground footwall damage is primarily controlled by large-scale slope failure mechanisms. Structurally controlled near-vertical planar shear failures in the upper footwall appear to be driven or facilitated by a step-path like curved shear failure situated in a zone starting ca 250 m, and intersecting the SLC about 100 m above the active mining level. To confirm this failure evolution a pilot underground measuring system has been designed and installed to monitor the expected rock mass displacements indicated by numerical models based on the above mechanisms. The systems consists of 50 m time domain reflectometry (TDR) coaxial cables installed in the footwall both across and away from an estimated damage extent boundary to monitor shear movements along natural joint surfaces. The TDR installation is combined with long extensometers to facilitate differentiation between diffuse shear and rock mass normal expansion. Additionally, a low-tech system of tape extensometer lines are installed in drifts and stopes running perpendicular to the ore strike to monitor horizontal displacements in the infrastructure as the footwall ‘slope’ loses confinement during mining advance when the mined out areas are replaced by caved rock from the hanging wall.CITATION:Svartsjaern, M, Saiang, D and Mäkitaavola, K, 2016. Underground footwall monitoring at the sublevel caving mine – Kiirunavaara, Sweden, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 773–780 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016

By De Jong HJ

During the forthcoming reline of No.3 Blast Furnace at the Newcastle Steelworks of the Broken Hill Proprietary Co. Ltd., facilities for the screening and weighing o' coke will be installed. The equipment is needed to optimise the sizing and the quantity of coke immediately prior to it being charged to the furnace. Up till the time the furnace was blown out in November, 1982, coke was charged to the furnace by volume, ie by skip fulls only. This resulted in variation to the ore/coke ratio charged due to bulk density changes and erratic skip filling levels. This in turn affects the heat balance of the furnace and causes fluctuations in hot metal chemistry and temperature. A coke rate penalty has to be paid also. Coke layer thickness in the furnace burden cannot be controlled closely, resulting in loss of control over gas distribution in the furnace. A grizzly type screen was incorporated in the skip loading chute but this ineffective method left a great amount of fines with the coke being charged to the furnace. Furnace permeability was therefore reduced, particularly in the hearth.

Jan 1, 1984

By M R. Drumond

The present work presents a new theoretical interpretation of the flotation process, which can also be applied to other mineral processing physical unit operations. This interpretation is part of the æModelo OperacionalÆ (Operational Model) that was developed as an attempt to establish a new theoretical basis for mineral processing unit operations. In the case of flotation, a book was edited by one of the authors in Brazil (Mass Concentration by Froth Flotation, in Portuguese and Spanish).The model establishes a link between a natural phenomenon (migration of particles from the pulp to the froth) and a real industrial phenomenon (mass concentration by operational interference). In this model, flotation is viewed as a phenomenon of macromolecular mass transport (of particles) between froth and slurry phases. Mechanisms for mass transport and particle flow between phases as a function of the hydrophobic potential in each phase are discussed. By establishing the so-called equilibrium distribution curve between phases it is possible to calculate macroscopic balances that are related to metallurgical results, which in turn are a function of the operational line that is required by the process itself. The number of cleaning stages and their efficiencies can also be determined by simple laboratory-scale experiments.

Jan 1, 2005