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By Wing G. Agnew, Paul L. Russell

"INTRODUCTION During the past few years, the development of millisecond-delay electric blasting caps has made possible a more flexible control over timing the firing of individual charges in blasting, The use of these caps has proved especially productive in quarry blasting and in numerous construction operations, but their use in underground mining is developing more slowly, although in some cases, as the solution to a difficult breaking problem, they have been well-received. When used in metal-mine drifts, the general objection is that the broken rock is thrown too far. In nearly all cases, however, use of millisecond-delay caps results in better fragmentation of the rock and a lower powder consumption per ton of rock broken.Prom 1949 to 1952, several series of experiments were conducted at the Mount Weather, Va., Experimental mine in conjunction with driving a 6- by 10-toot ventilation raise. 3/4/ The raise-round experiments in this 238-foot shaft, using millisecond-delay electric blasting caps in various drill-round patterns, are the subject of this report.SUMMARY AND CONCLUSIONSAll rounds discussed in this report were drilled in the hard rock (epidosite and greenstone) of the ventilation raise at Mount Weather Experimental mine (figs. 1 and 2). The drill-round patterns varied from 28- to 32 holes and from a V-cut design to one having no cut boles at all. With the exception of the first two rounds, all blasting was done with millisecond-delay electric blasting caps. In all rounds 60-percent ammonia gelatin dynamite was used and holes were stemmed with two cartridges of sand stemming."

Apr 1, 1953

By I Treloar, T Elenbaas, E Isaaks

Critical reviews of grade control generally focus on blasthole sampling and the estimation of ore control block model (OCM) grades, with little or no attention given to dig-line design. For example, the estimation of block grades is deemed important as estimation errors lead to the misclassification of ore types and subsequent dollar loss. Recognition of the problem that the underestimation of block grades may be more or less costly than overestimation led to the application of loss functions to the classification of individual OCM blocks by ore type. The method is known as maximum profit selection (MPS). However, each OCM block may be misclassified a second time by dig-line design. Although each block may initially be assigned an optimum ore type by MPS, OCM blocks are seldom individually mineable by ore type. As a result, a common solution to this problem is the design of mineable polygons or dig lines loosely constrained by a minimum mining width. All material within a properly designed dig-line polygon can be mined and sent to a single destination such as the mill, stockpile or waste dump. These dig lines are often designed manually or by simple variations of computer-generated contour lines. They are generally suboptimal and result in additional misclassifications of ore types. This paper provides an algorithm for constrained optimum dig-line design where dig-line misclassifications are also evaluated through loss functions or MPS. Constrained optimum dig-line designs minimise the dollars lost by dig line misclassifications for a given minimum mining width constraint. A case study is provided illustrating constrained optimum dig-line design and subsequent benefits.CITATION:Isaaks, E, Treloar, I and Elenbaas, T, 2014. Optimum dig lines for open pit grade control, in Proceedings Ninth International Mining Geology Conference 2014 , pp 425–432 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 18, 2014

By Richard F. McFarlin

INTRODUCTION Phosphate, one of the three major elements essential for all living plants and animals, has been mined in Florida for over a hundred years. Comprising roughly 80% of the U.S. total and 25% of the world's supply, the bulk of the 35 to 40 million annual ton output is converted to chemical fertilizers. With the exception of one mine in north Florida near the Georgia state line, today's mining activities are centered in the middle of the peninsula just east/southeast of Tampa in Polk, Hillsborough and Hardee counties. These "bone valley" deposits extend into Manatee and DeSoto counties, but there is little active mining there as yet. Because strip mining is, perhaps, the ultimate environmental insult and because Florida has a proactive environmental community, it is not surprising that the mining of phosphate rock has become increasingly controversial. This paper describes how phosphate rock is currently mined, the environmental problems presented by mining and processing the rock, and some mitigating technologies being employed to answer such problems. THE PHOSPHATE DEPOSIT AND ITS EXTRACTION Florida's phosphate deposits containing the calcium phosphate mineral, carbonate-fluorapatite, are between ten and fifteen million years old. The typical "bone valley" matrix is 15 feet thick and lies under 20 to 30 feet of overburden.(Fig. 1) Both the sandy overburden and matrix are unconsolidated and can be removed with standard 45 to 65 cubic yard, walking electric draglines. After removing the overburden, which is windrowed nearby, the dragline operator casts the matrix into a "slurry" pit where it is broken up by high pressure water monitors. Pit pumps (with boosters where necessary) transport the matrix slurry through 16 to 20 inch pipelines to a washer where the three components of the matrix (sand, clay and apatite) are separated. The phosphates contained in central Florida matrix are approximately half pebble/half flotation feed, through the pebble ratio decreases as the industry mines southward. Again, the matrix is roughly equally divided among sand, clays and apatite. The phosphatic clays, 75% of which are less than one micron in size, leave the washer as a 3 to 5% colloidal suspension. The impoundment of these waste clays is one of the environmental challenges to be discussed later. The sand (silica) is separated from the similarly sized phosphate particles by a double float process (fatty acid followed by a reverse mine or cleaner float), and the sand tailings are then pumped to disposal or reclamation sites. The apatite flotation product is combined with the screened phosphate pebbles and stacked to dewater before shipping.

Jan 1, 1992

By G. Sachs, Jr. Brown W. F., Ebert L. J.

It is a generally recognized fact that a steel quenched from the austenitic range and tempered to yield a structure composed of tempered martensite becomes increasingly more ductile with increasing tempering temperature. However, the conventional tensile test fails to yield any value that is sensitive to differences in behavior on tempering among the various low-alloy steels or among different heats of a particular steel. The notched impact bending test, on the other hand, has been recognized for a considerable length of time as capable of revealing differences in impact properties among various steels tempered to 'the same strength level (i.e., the same tensile strength in a conventional tensile test). However, this test is very complex and its significance has not been clarified up to the present time. Of its three characteristic features—the high speed, the notch, and the bending—the notch has been found by Davidenkov2 to be the most potent embrittling factor. Recent investigations on low-alloy steels have shown static notch-bar tensile tests to be more lucid and to evaluate the metal properties in much the same manner as impact tests. In an attempt to analyze the fundamental nature of this test, it was found possible to separate the effects of the two basic characteristics of the stress pattern introduced by the presence of a circumferential notch, the stress state and the high initial stress concentration at the notch bottom. The major conclusions drawn from these previous investigations were as follows: I. The magnitude of the triaxiality at fracture depends primarily on the initial notch depth. Its effect is a great reduction in the average ductility (contraction in area at fracture in the notched section) as compared with that observed in uniaxial tension. This triaxial stress state occurs at a point in the interior of the notch specimen. It determines the notch ductility if the notch is mild and/or the metal possesses a considerable ductility. On the contrary, brittle specimens probably fracture on the metal surface (at the notch bottom) at which point biaxial stress state is present. This biaxiality also considerably reduces the ductility of the metal. For small notch ductilities the notch strength and the notch ductility become very sensitive to embrittling factors. This is contrary to the behavior. of the ductility values derived from a conventional tensile test, which are comparatively insensitive.

Jan 1, 1947

By Ebert L. J., G. Sachs, Jr. Brown W. F.

It is a generally recognized fact that a steel quenched from the austenitic range and tempered to yield a structure composed of tempered martensite becomes increasingly more ductile with increasing tempering temperature. However, the conventional tensile test fails to yield any value that is sensitive to differences in behavior on tempering among the various low-alloy steels or among different heats of a particular steel. The notched impact bending test, on the other hand, has been recognized for a considerable length of time as capable of revealing differences in impact properties among various steels tempered to 'the same strength level (i.e., the same tensile strength in a conventional tensile test). However, this test is very complex and its significance has not been clarified up to the present time. Of its three characteristic features—the high speed, the notch, and the bending—the notch has been found by Davidenkov2 to be the most potent embrittling factor. Recent investigations on low-alloy steels have shown static notch-bar tensile tests to be more lucid and to evaluate the metal properties in much the same manner as impact tests. In an attempt to analyze the fundamental nature of this test, it was found possible to separate the effects of the two basic characteristics of the stress pattern introduced by the presence of a circumferential notch, the stress state and the high initial stress concentration at the notch bottom. The major conclusions drawn from these previous investigations were as follows: I. The magnitude of the triaxiality at fracture depends primarily on the initial notch depth. Its effect is a great reduction in the average ductility (contraction in area at fracture in the notched section) as compared with that observed in uniaxial tension. This triaxial stress state occurs at a point in the interior of the notch specimen. It determines the notch ductility if the notch is mild and/or the metal possesses a considerable ductility. On the contrary, brittle specimens probably fracture on the metal surface (at the notch bottom) at which point biaxial stress state is present. This biaxiality also considerably reduces the ductility of the metal. For small notch ductilities the notch strength and the notch ductility become very sensitive to embrittling factors. This is contrary to the behavior. of the ductility values derived from a conventional tensile test, which are comparatively insensitive.

Jan 1, 1947

By Giuliana A. Zelada

This paper reports on a study of 43 compression load tests (five fully instrumented) and ten pullout tests of auger cast-in-place (ACIP) piles in sand. Eight design methods reported in the literature were used to predict the measured pile capacities. Ultimate load capacities of the ACIP load tests were determined using two different definitions of ultimate capacity and compared with predictions. Load transfer behavior was determined from a study of the fully instrumented pile tests and compared to predicted values. Recommendations are made regarding the design approach most appropriate for these foundations and the best procedure for estimating ultimate capacity from pile load tests.

Jan 1, 2000

By M. C. Bétournay

Mine site infrastructure for storage and delivery of hydrogen must be standardized and acceptable for routine and safe industrial use before fuel cell technology is adopted as a major power source. An initial study reviewed possible options for production, storage and delivery of hydrogen for underground equipment. The high level review included assessment of economic, operational, technical, and health and safety risks. Following this review, a subsequent theoretical study to select typical minesite infrastructure and for construction at an operating research mine is outlined.

May 1, 2007

By Tim Neil, O&apos

Copper price rebounds A worldwide scramble for diminishing supplies of copper has pushed the red metal's price to more than $2.20/kg ($1 per lb), the highest level since 1980. The first nine months of 1987, world copper production was down 1% from a year ago, while consumption rose 7%. Demand has been fueled by economic strength in several countries, particularly Japan, South Korea, and Taiwan. The price has also been helped by interest rate declines that have prompted manufacturers to build up inventories. In addition, China is expected to be a major buyer in 1988 as it stocks up for its five-year industrialization plan. There appears to be no sign of any significant increase in copper supplies until later 1988 or early 1989. This means supplies probably will remain tight for some time even if world economic activity slows. Strikes and other production problems slowed world output of copper in 1987. As a result, copper supplies are the lowest they have been in 14 years. World inventories currently total 625 kt (690,000 st). That is not much above the 585 kt (645,000 st) level reached in 1974 when the price of copper rose to a record $3.84/kg ($1.74 per lb). This makes all the more interesting what Chile, the world's largest copper producer, has been doing lately. Chile has been the world's largest copper producer since 1982, when it overtook the United States. Table 1 shows Chile's copper grades, reserves, and production at its mines. Chile contains roughly one-fifth of the world's copper reserves. At present production rates, these reserves would last for more than 100 years. In 1986, Chile produced 1.4 Mt (1.5 million st) of fine copper, up 3%. Chile accounted for 17% of the world's copper output.

Jan 1, 1988

By Kazuyuki Kakegawa

"Sintered body of lead zirconate titanate (PZT) was prepared through the spark plasma sintering (SPS) technique. When the mixture of raw materials was calcined at 800°C and sintered at 900°C for 10 min. with SPS technique, the sintered body had almost theoretical density but was a mixture of lead titanate (PT) and PZT. When a calcining temperature of 1100°C was employed, sintered body had a single PZT phase but a lower density. High dielectric constant was obtained for a sample calcined at 800°C, sintered with SPS tecbriique and fired again in the air at 1100°C.IntroductionSpark plasma sintering (SPS) is a recently developed technique for sintering metals and ceramics1. The scheme of SPS is shown in Fig. 1. Sample is put in a graphite die in a vacuum chamber. It is pressed and pulse voltage is applied. The electric current heats the die. It is said that the pulse causes a spark-plasma between particles in the sample. Because of the spark- plasma, sample can be sintered at temperatures lower than that in the ordinary sintering by several hundred degrees of Celsius. The sintering period is also very short2Low sintering temperature and short sintering period in SPS technique cause low reaction rate. It is both disadvantage and advantage. In some cases, the reaction cannot be completed under the SPS conditions. If a mixture of several compositions having different properties can be sintered without a compositional change by the diffusion during the sintering, combination of properties, and therefore a design of property will be possible3. The low reaction rate of SPS may enable such combination."

Jan 1, 2000

By J Foster

The Lanfranchi Nickel mine is located approximately 48 km south of Kambalda in Western Australia. Mining first commenced in the mid-1970s under the previous owner, Western Mining Corporation Ltd. After two years on care and maintenance, Panoramic Resources (previously known as Sally Malay Mining Ltd) purchased the operation in November 2004. The site currently produces approximately 10?000 tonnes of nickel metal annually from approximately 430?000 tonnes of ore. The Ore is mined from four main orebodies; Deacon, Helmut South, Lanfranchi and Schmitz. The Helmut South orebody is situated approximately 790 m below surface and extends to a resource depth of 940 m. Horizontal thickness of the orebody ranges from 10 m to 45 m in width, with the average being 30 m. Previous mining at the Lanfranchi operations specifically the Helmut orebody, had a history of seismic activity. When Panoramic Resources acquired the operation it was decided that seismicity was going to be an issue and therefore a mining method needed to be selected that could effectively manage this potential risk. The dip of the orebody is quite shallow at 45º, with a 30º plunge to the south . The hanging-wall comprises of very weak ultramafics. Due to these factors the orebody posed several mining challenges. To safely and effectively mine the Helmut South orebody in the Lanfranchi Nickel mine, underhand cut and fill mining was selected as the most appropriate mining method. Paste fill comprising of cemented gold tailings was the backfill of choice. This paper is intended to convey some of the key learnings from the use of a mining method that has not been commonly practiced in Australia.

Mar 21, 2011

The multiphase reactive flow in process equipments, e.g., fluidized beds, involves complex coupling among flow, heat/mass transfer and reaction kinetics over scales ranging from molecules, sub-particle eddies up to equipments. Conventional two-fluid model (TFM) ignores the effects of meso-scale structures, and hence is not suitable for describing the intrinsically multi-scale heterogeneity within multiphase systems. To be consistent with the meso-scale structures, we proposed a structure-dependent multi-fluid model (SFM), and closed its drag force by invoking the stability condition defined in the energy-minimization multi-scale (EMMS) model. The SFM reduces to the conventional TFM if homogeneous distribution is assumed within each grid, and reverts to the original EMMS model if it is used to describe steady state, global behavior of a fluidized bed. The multi-scale computational fluid dynamics (CFD), which is characterized by coupling of TFM and EMMS drag coefficient, can then be viewed as a simplified solution of the SFM. We demonstrate that the SFM significantly improves the accuracy and at the same time enables higher efficiency by allowing for much coarser grid used than in TFM. Recent application of it in 3D, full-loop, transient simulation of industrial processes, e.g., fluid catalytic cracking, coal combustion, gasification, kaolinite clay calcinations, shows its advantages in helping design and troubleshoot multiphase process equipments. Keywords: meso-scale, SFM, multiscale, CFD, simulation, structure, EMMS, multiphase

Sep 1, 2012

By B. J. Jody

This paper describes a one-step thennal-treatment process, developed at Argonne National Laboratory, to remove all organic contaminants from the glass-fiber waste that is generated during the glass-fonning process. The recovcred glass fibers can then be recycled into new glass fibers or new glass products without adversely impacting the operations of the glass industry or the quality of its products. Our process involves shredding the waste to a manageable size and then processing it at temperatures below its melting point to remove the polymeric-coating layer. The glass industry generates more than 60,000 tons of glass-manufacturing waste annually in the glass-forming process alone. Recycling of this waste stream can keep this non-biodegradable waste out of the landfills. In addition, use of recycled glass lowers energy costs, by an average of about $5 per ton as compared to using virgin raw materials to make virgin glass fibers. This process will also reduce the emission of C02 by 17,400 tons a year and it will also reduce the emission of nitrogen oxides. Economic analysis of our process shows a potential payback of less than 2 years.

Jan 1, 2003

By C. D. Mishra

The recycling potential for any gainful utilisation of the waste products is dictated by their quality, especially the chemical characterisation. The methods commonly used for the analysis of fly ash e.g. wet chemical, AAS, ICP-AES etc are time consuming. The paper describes a rapid XRF-based method for routine analysis of fly ash which involves preparation of local standards, optimisation of sample preparation technique and use of a simple PC-based software to account for interelement effects. The matrix corrected values are found to agree well with the wet chemical data.

Jan 1, 1992

By Frank W. Sansom

This paper is one of a series of articles dealing with milling methods and costs at metal mines in the United States, written in accordance with an outline prepared by engineers of the United States Bureau of Mines and published by the bureau. The Netta mine and concentrator is the largest of the plants of the Eagle-Picher Lead Co, It is located in Ottawa County, in the extreme northeast corner of Oklahoma, and in what is known as the Tri-State zinc and lead mining district, embracing the northeast corner of Oklahoma, the southwest corner of Missouri, and the southeast corner of Kansas. A majority of the production comes from the Oklahoma section, with Kansas second, and Missouri third.

Jan 1, 1930

Underground mining has the potential to be one of the more dangerous industries in the United States and the world. Following recent disasters in Sago, WV; Darby, KY, and Crandall Canyon, UT, the mining industry was reminded of the need for a system to allow trapped miners awaiting rescue to survive following a disaster. To address this need, Battelle has developed a Mine Barrier Survival System (MBSS) that surpasses all current technologies and sets a new standard for keeping miners alive in the event of a fire, flood, explosion or collapse. The MBSS creates a temporary refuge for miners who cannot escape by using inflatable barriers and a post-fire atmosphere scrubbing system to define the ?safe haven? and remove CO2 and CO while replacing oxygen and providing heat. Current personal-protective systems in use are designed to provide breathing air for only a few hours when, realistically, it may take days for rescue teams to reach trapped miners. Battelle?s MBSS is designed to aid up to 15 miners for four days.

Jan 1, 2008

By R. F. Holub, E. Franco, T. B. Borak, K. J. Schiager

INTRODUCTION According to a recent analysis by Schiager et al.(1981), inhalation of radon daughters continues to contribute the largest risk from exposure to radioactive materials in underground uranium mines. The American National Standards Institute (1973) has published a concensus standard for radiation protection in uranium mines. According to this standard, a monitoring system for radon daughters must be capable of measuring the annual accumulated exposure in working level months within an uncertainty interval of 50% at the 95% confidence level. To satisfy this criterion, uranium mining companies in the USA have adopted a procedure consisting of single or periodic measurements of the air concentration of radon daughters. All of the measurement methods sample a known volume of air through a filter. After a specified delay interval, the filter is analyzed for radioactivity. The exposure rate from total airborne potential alpha energy in units of working level (WL) is estimated from the measured activity on the filter. For control purposes, this measurement establishes whether or not a work area is safe for occupancy by miners. For routine dosimetry, the concentration estimated from the grab sample is assigned to the appropriate work area. It is then assumed that the concentration of radon progeny is equal to the average of the two measurements at the beginning and end of the specified interval. Meanwhile, the work locations and exposure times of each miner are recorded on the daily time card and verified by the shift supervisor. Most cards provide space for 15-minute time intervals. The exposure in working-level hours (WL-hr) for each shift is obtained by summing the time spent in a given location multiplied by the exposure rate (WL) assigned to that location.The exposure of an individual miner to radon progeny is summarized on a weekly basis and commonly reviewed each month. These values are then accumulated in units of working level- months (WLM) during the calendar year. EVALUATION TECHNIQUE AND RESULTS The evaluation of grab sampling methods for monitoring exposure to radon progeny is presented in two sections. The first section examines each method individually according to accuracy, precision, and uncertainties attributed to human errors. The second section includes an analysis of procedures which are common to all measurement methods. The ultimate goal is to determine whether a specific measurement method, or grab sampling in general, can fulfill the criterion that the annual accumulated exposure of each miner be known to within ± 50% at the 95% confidence level. In addition it is essential that the above criterion be satisfied at an acceptable cost to the mining company. The results of each testing procedure are presented in terms of the contribution of the relative uncertainty in the estimates of cumulative WLM. Each of these are then combined to form a figure of merit for the complete monitoring system. The method of combining statistical uncertainties and interpretation of the results is subject to some debate. However, since all of the tested parameters enter into the final computation of WLM multiplicatively, we have used the commonly accepted procedure of combining the squares of the relative uncertainties. We have also assumed that each of the uncertainties is derived from independent random variables and the covariance of any two variables is zero. A description of each grab sampling measuring method is given in Table 1. Also included is the equation used to compute WL from the measured quantities. Accuracy (U A) Our definition of accuracy is based on the uncertainty inherent in any method which attempts to estimate potential alpha energy without unambiguously determining the concentration of the three unknown radioisotopes, RaA, RaB and RaC. (For simplicity, the RaC and RaC' nuclides are referred to in combination as RaC.) The inherent uncertainty strongly depends on the mixture of these three components at the time of measurement.

Jan 1, 1981

By C E. Gawith

The Waihi underground mining operation currently mines epithermal veins ranging from two metres up to ten metres in width using a ‘Modified Avoca’ stoping method. This method consists of longhole bench stope mining at 15m to 18m sublevel spacing. The bulk of production drilling and blasting is top down with the extraction length of a typical stope panel nominally being 18m to 20m along strike. A stope panel is usually taken in two blasting cycles followed by the remote bogging of the stope ore. The panel is backfilled in-cycle using unconsolidated waste rock material which is tipped into the void from the upper level using a Caterpillar R2900 or R1700 sized underground loader. The underground mine first commenced production in 2008 and up until 2016, as was typical of many open stope mining operations at the time, an underground loader was used to place a waste rock bund at the top brow of the open stope. This rock bund was the principal means of protecting people and equipment from this hazard. In 2016 there was a fatality at the mine when a loader that had been tasked with building this rock bund fell into the stope void. Immediately post incident all mining and processing operations were suspended for a week to allow colleagues to grieve, to observe cultural protocols and to begin the investigation into this event. During this period senior operators, supervisory and technical staff developed an innovative system which used steel bollards to protect this open edge. With development over the following months this bollard system has eliminated this hazard from the Waihi underground mine.

Mar 15, 2021

By V. Borissenkob, L. D. Ropera, M. J. Danielc

Eccentric High Pressure Centrifugal Comminution (eHPCC) is a novel comminution device that provides a single step to achieve a desired target grind size. The mechanisms within the unit mirror particle breakage that is akin to high-pressure-grinding rolls (HPGR) and high-intensity-attrition with air classification. eHPCC receives lump and can produce a product P80 such that the progeny are at or less than natural grain size of the feed. The unit operates in either closed or open-circuit. The grinding zone is dry and free of steel grinding media. This paper presents evidence of eHPCC-2 capable of selective and preferential mineral liberation. A recent test campaign on a blend of underground quartzgold- vein ore was completed. The feed to eHPCC-2 trial was SAG mill trommel oversize (low-grade pebble rejects). eHPCC-2 operated open-circuit with air-fluidisation in the rotating-grinding-chamber, classifying particles based on terminal velocity of the air-swept product particles, displaying size and mineral composition classification taking place at the same time. There is an opportunity for future industrial research targeting eHPCC as an advanced integrated comminution and classification technology. Keywords: High pressure, high intensity, compression, attrition, comminution, classification

Jan 1, 2020

Roof bolting has been widely used in hard rock underground mines for its many advantages. In this paper, the applicability of roof bolts to mine openings in frozen ground and effects of bolts on the stability of the openings are presented. A comprehensive finite element program ABAQUS, was used to analyze the temperature distribution, the stress field, and the displacement field around a mine opening. Openings supported by roof bolts and unsupported were analyzed. The thermal and mechanical processes were treated as a quasi-coupled problem, with the thermal analyses being conducted first, followed by the mechanical analyses in each case. For the opening under consideration with a roof area of 15 m2, five to seven roof bolts are appropriate. When the number of bolts are less than five, the displacement within the ground can be improved only locally.

Jan 1, 1995

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