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By S. K. Mishra

The history of the Australian coal industry is almost as old as the history of Australia after new settlement. The first cargo of coal was exported from Newcastle to India in 1799. Although coal has been mined in Australia for almost two centuries, there are indications that the first coal was processed in 1877 in the New-castle area, nearly a half-century after the first coal processing in Europe. Processing on any significant scale began only in 1944 at the Newcastle Steel Works. In Queensland, however, coal processing did not begin until the second quarter of this century. The well-known dramatic growth of mining and processing of Australian coal since the mid-sixties has primarily been due to the expansion of the export market in Japan and other Asian and European countries. Black coal occurs in most of the major sedimentary basins in Australia. The better known deposits of commercial importance are in Sydney, Bowen, Oakland and Gailiee Basins (Figure 1). Apart from Collie Basin in Western Australia, other coal deposits of the Arckinga Basin in South Australia, the Bonaparte Basin in Northern Territory and Tasmania Basin, Tasmania, are basically unexploited. Sydney Basin of New South Wales and Bowen Basin of Queensland remain the major producers of black coal in Australia.

Jan 1, 1985

By R. J. Mirville

Flocculation is an important step in the treatment of coal preparation waste waters. The effects of polymer charge and molecular weight and solution pH and ionic strength on the adsorption and flocculation characteristics of a series of polymeric flocculants on some typical waste water constituents have been studied experimentally. The results indicate that the nature and extent of flocculation are strongly system-dependent. For the case of coal, both adsorption and flocculation behavior can be correlated with the polymer conformation in solution. For the mineral constituents, on the other hand, adsorption and flocculation depend also on the surface properties of the solid.

Jan 1, 1979

By A. T. Visser

The expansion of surface mining has led to the development of very large off-highway trucks currently capable of hauling payloads in excess of 290t. Mine haul roads have historically been designed empirically, relying heavily on local experience; the trend in increasing truck size will thus render the current pavement systems inadequate. Since truck haulage costs can account for 10%-50% of the total costs incurred by a surface mine, not only would the maintenance costs of existing roads of inadequate design increase, vehicle operating and maintenance costs would also increase prohibitively. There is thus a need for improved design technologies encompassing the construction and management techniques of mine haul roads, appropriate for the wheel loads of vehicles now in use. The aim of the paper is to present the different components of a mine haul road design and management system and to demonstrate the value of its application through case studies. The system has been applied at several South-African surface mines and distinct benefits have been derived. The improved structural design of a new road resulted in a 29% saving in construction costs and also provided better service, whilst the optimal selection and management of wearing-course materials provided better functionality at lower total transportation operating costs.

Jan 1, 1999

By William M. Warren

Recent sinkhole activity in carbonate terranes of Alabama has resulted in damage to transportation lines, structures, utility lines, and property in general. Active sinkhole development and subsidence has been well documented in Alabama, Florida, Pennsylvania, and South Africa. An excellent example of the potential danger resulting from this activity is the spectacular collapse under a South African gold ore crusher plant that resulted in the death of 29 men in 1962. Due to the severity of the problem in certain localized areas in Alabama, a detailed investigation is being undertaken by the Geological Survey of Alabama and the U. S. Geological Survey. One of the most promising techniques is the application of aerial photography and thermal infrared imagery. Conventional and modern remote-sensing techniques are being used in Alabama to locate and define areas of sinkhole activity and to determine which areas are most prone to its occurrence.

Jan 1, 1975

By P. Randall, J. Ickes, S. Chattopadhyay

This study was undertaken as a part of developing treatment alternatives for waste materials, primarily waste rock and roaster tailings, from sites contaminated with mercury (Hg) mining wastes. Leaching profiles of waste rock over a range of different pH and oxidation-reduction (Eh) conditions were performed. Since iron (Fe) is present in large quantities in mining wastes, the effect of Fe on leaching of Hg was also determined. Results show that the Hg concentration increased as the pH of the suspension increased until a pH value of 10.65. Thereafter, the Hg concentration decreased sharply. The Hg concentration, in presence of Fe decreased significantly (1/10thto 1/100th). Kinetics results show a relatively slow rate of leaching in presence of ferric ions. The results obtained from the present study are being used to predict the fate and stability of Hg present in the environment, and will therefore, be useful in dictating the nature and suitability of any remediation treatment.

Jan 1, 2003

By W. W. Wen

The objective of this investigation was to study the electrokinetic behavior of oxidized coals to indicate the feasibility of separating oxidized coals from ash materials and pyrite by an mine flotation process. The effects of soluble inorganic electrolyte products from coal oxidation were also examined. The electrokinetic behavior of coals of different rank and degree of oxidation were studied by an electrophoreses technique. Generally, increasing the oxidation increases the negative value of the zeta-potential and decreases the isoelectric point (IEP). Flotation results indicate that the flotation recovery increases with decreasing pH values, and that with ferrous sulfate at pH 6.5 the flotation recovery is optimum while the pyrite was depressed. Comparing the flotation results with the electrokinetic results verified that in the presence of ferrous sulfate, the maximum hydrophobicity of oxidized coal with dodecylamine and frother is near its IEP which is around pH 6.5 to 7.0.

Jan 1, 1976

By L. Yuan, A. Smith

"NIOSH’s mine fire simulation program, MFIRE, is widely accepted as a standard for assessing and predicting the impact of a fire on the mine ventilation system and the spread of fire contaminants in coal and metal/nonmetal mines, which has been used by U.S. and international companies to simulate fires for planning and response purposes. MFIRE is a dynamic, transient-state, mine ventilation network simulation program that performs normal planning calculations. It can also be used to analyze ventilation networks under thermal and mechanical influence such as changes in ventilation parameters, external influences such as changes in temperature, and internal influences such as a fire. The program output can be used to analyze the effects of these influences on the ventilation system. Since its original development by Michigan Technological University for the Bureau of Mines in the 1970s, several updates have been released over the years. In 2012, NIOSH completed a major redesign and restructuring of the program with the release of MFIRE 3.0. MFIRE’s outdated FORTRAN programming language was replaced with an object-oriented C++ language and packaged into a dynamic link library (DLL). However, the MFIRE 3.0 release made no attempt to change or improve the fire modeling algorithms inherited from its previous version, MFIRE 2.20. This paper reports on improvements that have been made to the fire modeling capabilities of MFIRE 3.0 since its release. These improvements include the addition of fire source models of the t-squared fire and heat release rate curve data file, the addition of a moving fire source for conveyor belt fire simulations, improvement of the fire location algorithm, and the identification and prediction of smoke rollback phenomena. All the improvements discussed in this paper will be termed as MFIRE 3.1 and released by NIOSH in the near future INTRODUCTION The MFIRE program, originally developed in the 1970s by the U.S. Bureau of Mines (now the Office of Mine Safety and Health Research (OMSHR) at the National Institute for Occupational Safety and Health (NIOSH)) and Michigan Technological University, has been a mainstay in modeling of the mutual influences between a fire and the mine ventilation system during an underground mine fire. Several update versions have been released over the years, including MFIRE 1.27, 1.29, 1.30, 2.0/2.01, 2.10, and 2.20. The MFIRE program logically comprises four parts: (1) a conventional network calculation, where it performs the basic network balancing without considering heat or mass transfer; (2) a temperature calculation to establish the reference temperature distribution before a non-steady state (transient-state) simulation; (3) a transient-state simulation that follows changes in ventilation step-by-step to produce a continuous description of the temperature distribution, smoke, and contaminant spread through the ventilation system during a fire event; and (4) a quasi-equilibrium simulation to predict the state of the ventilation system after a relatively long period of time (defaulted as 5 hours in MFIRE) as the fire reaches a quasi-steady state. In summary, MFIRE is a computer simulation program that performs normal ventilation network planning calculations and dynamic transient-state simulation of ventilation networks under a variety of conditions including the influence of natural ventilation, fans, fires, or any combination of these (Chang et al., 1990)."

Jan 1, 2016

By Carl G. Hogberg

To the casual observer there is probably no simpler appearing operation in the mineral industry than a limestone quarry. It might seem that the entire operation - from the drilling, blasting, and loading of the stone, its transport to a mill for crushing, washing and sizing, to loading into boats, railroad cars, or trucks for delivery to customers - can be comprehended at a glance. But no first impression could be more misleading. It is one of the purposes of this presentation to sketch the broad outlines of the extensive limestone industry in the United States which, measured in terms of volume of production since the end of World War II, has grown almost 300 per cent to a national level in 1957 of 384 million net tons.1 Another objective will be to discuss in more detail this industry in the State of Michigan. For this latter purpose, the Michigan Limestone Division plants at Calcite and at Cedarville will be used as examples. These descriptions will, we believe, show that the limestone business has complexities to match its size.

Jan 1, 1962

By J. Hanna, J. Jung, I. J. Anazia

Electrothermal production of elemental phosphorous (P4) generates substantial amounts of highly toxic phossy water and sludges. Because of their high phosphorous contents and lack of reliable processing technology, large tonnages of these hazardous wastes have accumulated from current and past operations. To date, the sludges are stored in ponds or above ground in special containment vessels at various locations around the country. Two samples of electric furnace phosphorous sludge (obtained from the Tennessee Valley Authority (TVA) in Muscle Shoals, AL and FMC Corporation in Pocatello, ID) were investigated using physical separation techniques to recover the elemental phosphorous and to render the residual material nonhazardous. The techniques tested include sizing, froth flotation and dispersed air oxidation. Flotation and sizing produced concentrates containing 69% - 90%P4, with recovery in the range of 60% - 87%. Dispersed air oxidation of the fines and tailings products was effective in the removal of an additional 10% - 15% of the P4 in the sludge.

Jan 1, 1993

By Heriban Soto

Primary amines are strong collectors of apatites and they can selectively float sedimentary phosphates from calcareous ores. However, secondary and tertiary amines and tri-methyl ammonium salts were ineffective. The use of kerosene reduces the collector requirements because hydrocarbon is coadsorbed along with the amine on the mineral surface. The heats of reaction of aminium ions with such anions as phosphate, carbonate and sulfate, were measured in a microcalorimeter and discussed in terms of the solubility of the precipitates formed and their relationship with the adsorption mechanism of primary amines on phosphate minerals. Experimental evidence supports a view that primary amines are adsorbed through chemical interaction between the aminium and phosphate ions at the mineral surface.

Jan 1, 1984

By Talaat A. B. Lawendy, Guerry H. McClellan

Batch laboratory-scale flotation tests were conducted on some carbonate-containing sedimentary phosphate ores to obtain concentrates suitable for chemical processing. The phosphate samples tested contained both calcite and dolomite. The calcitic phosphate ore was from Israel and the dolomitic ores were from China, India, Egypt, and Pakistan. The phosphate sample from Pakistan also contained a significant amount of silica. Several flotation techniques that have been proposed for upgrading carbonate -containing phosphate ores and concentrates were tested. Two processes gave better results compared with the other processes tested: a modified IMC anionic process and a. phosphoric acid depression process. At the best flotation conditions tested, phosphate concentrates assaying 32.5%-36.5% P205, 44.0%-51.0% CaO, and 0.6%-1.0% MgO were obtained at 69%-76% overall P205 recoveries from run-of-mine (ROM) ores containing 16.0%-25.0% P205, 40.0%-52.0% CaO, and 1.0%-11.6% MgO.

Jan 1, 1993

By R. F. Pearce

The paper in based on laboratory and pilot plant investigations which led to the development of a series of processes for production of high purity metals from nickel-copper-cobalt mattes of a wide range of composition. The selection of ammonia pressure leaching or acid pressure leaching for various mattes is discussed. This is followed by discussion of various methods for leach solution purification and metal separation. The metals of high purity are produced by reduction with hydrogen from purified solutions. Examples are given of complete flowsheets for a number of mattes. Drawing upon those examples, general principles for the selection of the most suitable treatment process for particular material are outlined. Under certain circumstances, those processes may replace part of the pyrometallurgical treatment and all of the electrolytic refining steps which are customary in the treatment of nickel-copper-cobalt mattes.

Jan 1, 1959

By B. R. Waples

Sometimes a project is born easy, sometimes a project has a hard time being born. Probably every project can say we had tremendous problems to overcome and they would all be right. This project had all things happen to it. Born easy, a hard time after being born, there were tremendous problems to overcome but here it is, a major copper project, a major operation, and a new source of copper for Mexico and the world market. La caridad is an ore body located in the state of Sonora in northwestern Mexico. It is at an elevation of 1,200 meters above sea level and the nearest major city is Hermosillo, the capital city of Sonora. Hermosillo is located approximately 270 kms (about 168 miles) to the southwest of la caridad. North of la caridad, 130 kms (about 81 miles), is the U.S. border town of Douglas, Arizona, which has a sister city on the Mexican side of the border -Agua Prieta.

Jan 1, 1979

By D. Stone, R. Pakalnis, B. Seymour

Most mines using backfill routinely carry out quality control (QC) testing of backfill mixes with unconfined compression test cylinders. This is true for paste operations, hydraulic fill operations, and cemented rockfill operations worldwide. However, a large number of mines simply do not prepare and test the backfill samples in a consistent manner and fail to properly analyze or interpret the test results. Conventional concrete test criteria are not applicable given the differences in the mix constituents and the fact that a large number of the backfill QC test cylinders typically fail to achieve the target strength. Standard industry practices are needed to improve the analysis and interpretation of QC test results and to provide a better means of determining if the backfill is achieving its intended target strength. This paper will attempt to lay the groundwork for standard industry practices and will address several key questions raised by operators, such as how laboratory test data compares to in-stope performance, and how many QC tests should fail before a red flag should be raised. INTRODUCTION Although paste backfill is widely used throughout the mining industry, cemented rockfill (CRF) remains as the backfill of choice in places like Nevada where very poor ground conditions mandate the use of underhand mining. The safety and stability of undercut entries beneath backfill is a function of the quality of the cemented fill, which is typically assured through routine quality control testing. In the absence of a backfill quality standard, most mines rely on concrete standards to design and interpret their backfill quality control programs. In the USA, these standards are developed and published by organizations such as the American Society of Testing and Materials (ASTM) and the American Concrete Institute (ACI). However, the application of these standards to cemented backfills can be very difficult owing to the high degree of variability in backfill mixes and the resulting variability in test results. To illustrate, the concrete industry standards denote that an acceptable (passing) test result occurs when the average strength of the duplicate test cylinders exceeds the target strength and no single test result deviates from the target strength by more than 10% (ACI 214R). Remedial measures must be taken if three consecutive tests fail. With mine backfills, it is safe to say that these criteria are rarely achievable. A further, more dramatic example is the accepted coefficient of variation for concrete QC cylinders. According to ACI 214R, the coefficient of variation (CV) should range from below 3% (excellent rating) to above 6% (poor rating). However, if this analysis method is applied to several backfill QC databases from operating mines, coefficients of variation are calculated on the order of 50%. Clearly, there is a need to develop better industry practices for assessing the quality of mine backfill and to assist mine operators with the interpretation of test results and the development of action plans in the event of undesired outcomes.

Jan 1, 2019

By Robert M. Stevens

INTRODUCTION Truck haulage of muck in underground mining operations has been in use around the world for many years. From the late 1.950s into the early 1970s, haulage by means of load-haul-dump (LHD) units challenged haulage by means of trucks, and, to a large degree, LHD units dominated the trackless concept of moving mate¬rial underground. However, beginning in 1975, the emphasis began to shift back to the use of trucks, with LHD units assuming the secondary role of loading the trucks. Over the years, several manufacturers have been offering trucks with capacities ranging from 4.5 to 45.4 t (5 to 50 st) and with features that specifically address the constraints of underground operation. TRUCK TYPES Three different types of trucks are available for use underground, namely the tip dumpers, telescoping dumpers, and push-plate dumpers. The following para¬graphs describe each of these three types. Tip Dumpers Tip dumpers are designed to lift the truck box, using gravity to dump the load out the rear of the box. The primary application of these trucks is in the haulage of muck to the surface. However, these trucks also may be utilized underground, if a chamber is provided to accommodate the raised truck box at the dumping point. Fig. 1 illustrates a four-wheel drive truck with a 10.0-t (11-st) capacity; this truck is only 1.9 m (6.25 ft) wide and is designed specifically to follow narrow and winding vein courses. Fig. 2 illustrates a two-wheel drive tip dumper with a 24.5-t (27-st) capacity; this truck operates on 12% grades with firm dry surfaces and in modes both loaded-up and loaded¬down. Telescoping Dumpers Telescoping dumpers specifically address the constraints imposed by low back heights, and they represent an optimization of compressed envelope size with maximum payload capacity. Figs. 3 and 4 illus¬trate a typical telescoping dumper with a capacity of 18.1 t (20 st). Fig. 3 shows the box in position for carrying the load, and Fig. 4 shows the same truck with the box in the full-dumping position. Fig. 5 illustrates the loading and dumping cycles. Fig. 6 illustrates an approximation of how an LHD unit with an eject-o-dump (EOD) bucket would be used to load over the rear of this type of truck. Fig. 6 shows a low-profile telescoping truck in the first stage of loading, with a 6.8-t (7.5-st) LHD unit equipped with an EOD bucket. A loader with a smaller capacity and a narrower bucket would allow the bucket to fit inside the truck box and would achieve an even lower loading height than that illustrated. Gen erally, an LHD unit equipped with an EOD bucket and teamed with a telescoping dump truck can be used with a lower back height than could any other front-end loader and truck team. Push-Plate Dumpers Push-plate dumpers are similar to telescoping Bump¬ers, except that they do not have the two-stage box. Instead, push-plate dumpers utilize a single push plate to discharge the load out the rear of the box. Gen erally, these units are limited to payload capacities of about 9 t (10 st). Fig. 7 illustrates a typical push-plate dumper.

Jan 1, 1982

By Xia Guangxiang, R. Y. Wan, Yang Hanlin, M. C. Fuerstenau

The mechanism and kinetics of pretreatment on pyritic gold concentrate with the catalytic oxidation acid leach (COAL) process were studied in the temperature range of 80-100°C and oxygen pressure range of 230-1013 kPa, with sodium lignosulfonate (SL) and mixed medium of sulfuric acid (11 g/[ ]) and nitric acid (5.5 g/ [ ]). The results show that nitric acid acts as a catalyst, the surfactant changes the hydro-phobicity of resulting elemental sulfur to favor the oxidation of pyrite, and the oxidation-determining step would be the chemical reaction between nitric acid and pyrite. When about 80% of pyrite contained in the concentrate is oxidized, the gold cyanidation extraction exceeds 95 % . The nitric acid in solution can be recycled and used again.

Jan 1, 1993

BP Minerals America (BPMA) has completed its $400 million modernization of Kennecott Utah Copper's Bingham Canyon mine. Modernized facilities include the in-pit crusher, conveyor, ore storage area, concentrator, ore pipe-line, and tailings pipeline. With modernization, Kennecott will achieve substantial cost reductions. Bingham Canyon is now one of the lowest cost copper producers in the world. And the mine will continue to provide jobs and income in Utah for another 30 years or more. The Modernization Project began in February 1986. The work took 33 months. In late 1987, the construction force peaked at 1100 workers. Kennecott Utah Copper now employs a permanent work force of 2100 employees. Less equipment and more efficiency The modernization work has increased the scale of the facilities that transport and process copper ore. A large, continuous transport system has replaced multiple haulage units. Fewer but larger pieces of equipment have replaced many crushing, grinding, and flotation units. The result is less maintenance, fewer employees, and improved metal recoveries. About one-third of the $400 million modernization project is directed at improving ore movement by eliminating rail transport. The new concentrator accounts for the remaining two-thirds of the modernization investment. The modernized plant was started up in stages. In February, ore was fed to the crusher. By the end of March, the three grinding lines were operating and the flotation circuit was producing copper concentrate. In July, the design through-put rate of 70 kt/d (77,000 stpd) was reached. October saw the achievement of targeted metal recoveries that corresponded to the fully modernized operation. It is estimated that Bingham Canyon will annually produce 181 kt (200,000 st) of refined copper, 9.3 t (300,000 oz) of gold, 71 t (2.3 million oz) of silver, and 5.4 kt (12 million lb) of molybdenite. Current Bingham Canyon production is about 97 kt/d (107,000 stpd) and a similar amount of overburden. Some 70 kt/d (77,000 stpd) of ore are processed through the in-pit crusher and 27 kt/d (30,000 stpd) of ore are transported by rail. The Bingham Canyon mine has yielded more than 11 Mt (12 million st) of refined copper. In 1863, precious metals - silver-bearing lead - were first discovered in Bingham Canyon. What was once a mountain is now a huge open pit. In fact, the mine is the world's largest man made excavation. The pit is some 4 km (2.5 miles) wide and 0.8 km (0.5 miles) deep. Since the pit began operating in 1906, more than 4.5 Gt (5 billion st) of material have been removed. Modernized facilities Electric 21 m3 (27 cu yd) or 23 m3 (30 cu yd) shovels are used to load the haulage trucks. Both 154 t (170 st) and 181 t (200 st) trucks are used to haul the ore to a mobile, in-pit crusher. The crusher replaces three crushers located at the old concentrators. There are two dumping locations at the new crusher, which can handle two trucks a minute. The 1.5 x 2.7 m (60 x 109 in.), 745 kW (1000 hp) Allis-Chalmers gyratory crusher reduces the ore to 254 mm (10 in.) size or less and delivers it to the conveyor. The crusher and conveyor systems are capable of handling 9 kt/h (10,000 stph) of ore.

Jan 11, 1988

By J. R. Hunter

The present is rooted in the past. This is particularly true in the case of mining operations. Coal mining in the United Kingdom is considered an old Industry comprising today 246 Coal Mines many of which have worked coal reserves in their locality for over 100 years. More than half the current producing mines lie in the Yorkshire and Midlands coalfields. They are in the oldest age group and produce 70% of the Industry's output. Some 50% of the remaining U.K. workable reserves amounting to approximately 98,000 million tons are still available to these regions. Fig. 1 Over the past two decades a gradual erosion of demand for coal in the U.K. has virtually halved the Industry's capacity. The number of Mines has reduced by 53% during the past 10 years. Productivity has increased by some 42% mainly through improvements in the central coalfields which has enabled costs to remain broadly competitive with oil.

Jan 1, 1976

By Gordon Clark, Basar Arioglu, Zlatanic. Sanja, Nasri Munfah, Togan. Tolga

"The 13.2 m external diameter double deck Istanbul Strait Road Tube Crossing is one of the most challenging underwater tunnel projects in the world. Subjected to 11 bars of hydrostatic pressure, variable geology, and being in a very active seismic area, the execution of the project required innovative solutions, sophisticated construction techniques, and prudent risk management approach. The 5 km tunnel consists of 3.4 km TBM bored tunnel, 1 km twin NATM tunnels, and cut and cover sections. The project is being delivered in a Build-Operate-Transfer (BOT) contract in a public-private partnership. The technical challenges, coupled with financial and commercial risks required the implementation of sophisticated risk management tactics including the provision of Independent Design Verification. Presently, the project is in its advanced stages of construction with a recent successful TBM breakthrough on August 22nd 2015. The paper presents the technical challenges of the project from the design, construction, and risk management aspects and provides the status of the construction. PROJECT BACKGROUND The $1.245 billion Istanbul Strait Road Tunnel Crossing project, dubbed the Eurasia Tunnel, is one of several major infrastructure projects being implemented in the Republic of Turkey in a public-private partnership approach. The project provides a direct and easy connection between the Anatolian (Asian) side of Istanbul and the heart of its historical district on the European side across the Bosphorus Strait which connects the Black Sea with the Sea of Marmara. Upon its completion in 2017, the project will improve connections to a wide network of motorways on both sides, increase capacity across the Bosphorus by 100,000 vehicles a day, reduce congestion and save motorists an average of 45 minutes of commuting time in each direction, and bringing significant economic benefits to the city and the region; it will ease traffic across the strait, reduce congestion, decrease pollutants and emission while maintaining the historical silhouette of the city. Figure 1 provides a general view of the project location and alignment. The overall project consists of three parts totaling 14.6km. Parts 1 and 3 consist mainly of widening existing motorways, providing connections to existing roads, reconstructing local bridges and underpasses on the European and Anatolian sides respectively. Part 2 is the tunnel crossing across the Bosphorus; it is the most complex part of the project with the greatest challenges and risks. Part 2 is 5.4 km long consisting of 3.4 km of 13.2m external diameter TBM bored tunnel under the Bosphorus Strait, 1-km of twin NATM tunnels on the Asian side, cut and cover transition boxes in the Asian and the European sides and depressed approaches on both sides. In addition, this segment includes the toll plazas, ventilation buildings, and the tunnel control and maintenance facilities."

Jan 1, 2016

By Charles T. Staniek

Following a dragline relocation at Agrico's Fort Green phosphate mine in 1986, the recovery plant experienced a significant decrease in flotation performance. Through weeks of trial and error testing, we discovered we could restore recoveries by conditioning feed at pH's higher than we had used historically. The higher pH's essentially increased flotation cell kinetics, although we still do not understand the exact mechanism at work. Armed with new-found appreciation of pH modification, we initiated a program to determine the optimum modifier for us, from both a performance and environmental perspective. Environmental considerations were important because ammonia, the standard industry modifier, is a plant nutrient and higher usage rates could affect the quality, of our recycle process water. This paper describes the research and development work spanning a 3 year period that culminated in making a major change in our reagent suite when we replaced ammonia with soda ash on a permanent full-time basis.

Jan 1, 1993