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By C. J. Pritchard, R. E. Gundersen

I commend the authors on examining an interesting method of shaft lining and attempting to improve safety in a dangerous environment by remotely installing the shaft lining. It is important, though, to not trade off shaft lining safety for long-term mine safety by lining an intake shaft with combustible material. In many operations, the mine utilities are installed next to the lining of the intake shaft, supplying a possible ignition source if damaged or overloaded. Although the polyurethane studied in the paper is of relatively high density, no mention is made of the flame retardency of the samples. When exposed to heat, small amounts of polyurethane will melt and bum but are extinguished when the heat source is removed. If large surface areas of combustible polyurethane - a tunnel or mine shaft- are exposed to ignition heat sources that are not removed or extinguished early, an uncontrollable, fuel rich fire may result with high carbon monoxide levels, low oxygen concentrations, and greatly increased burning rates that can have devastating effects on the ventilation system and miners' safety. Although these heavy polyurethanes lend themselves to a "pouring" application around the annulus of the shaft liners and is a continuous process, past experience with atomizing and frothing polyurethane spray systems indicates a high degree of maintenance and operator attention to the nozzle1 delivery system. This may require more miner's time to be spent in the shaft maintaining the foam system, negating some of the safety advantages of the system. If polyurethane foams are to be used in mines, they must be of the highest quality available in terms of flame spread, heat emitted while curing, and contaminants emitted while curing and spraying. In addition, employees must be given adequate training and be thoroughly informed in all aspects of polyurethane foam safety procedures. As witnessed by our fellow miners in South Africa, we can accept no less.

Jan 1, 1991

By Hang Yang, Linneas W. Laage

A major concern of mine fires is the heat generated ventilation disturbances which can move products of combustion (POC) through unexpected passageways. Fire emergency planning requires simulation of the interaction of the fire and ventilation system to predict the state of the ventilation system and the subsequent distribution of temperatures and POC. Several computer models were developed by the U.S. Bureau of Mines (USBM) to perform this simulation. The most recent, MFIRE, simulates a mine's ventilation system and its response to altered ventilation parameters such as the development of new mine workings or changes in ventilation control structures, external influence such as varying outside temperatures, and internal influences such as fires. Extensive output allows quantitative analysis of the effects of the proposed alteration to the ventilation system. This paper describes recent USBM research to validate MFIRE's calculation of temperature distribution in an airway due to a mine fire, as temperatures are the most significant source of ventilation disturbances. Fire tests were conducted at the Waldo Mine near Magdalena, NM. From these experiments, temperature profiles were developed as functions of time and distance from the fire and compared with simulations from MFIRE.

Jan 1, 1995

By David J. Dunn

Materials related problems at mines can be big ones - like a newly completed 1 -m (3.5ft) water pipeline burst during initial filling, grinding ball consumption at double the typical norms on startup of a new semiautogenous grinding (SAG) mill circuit or 30 open-pit haul trucks in the shop for new bed liners after less than a year’s service. Materials engineering provides solutions to these kinds of equipment problems by selecting good materials and designs, by trouble-shooting service failures and by maintaining high quality in vendor-furnished materials and components. Materials with good properties, correctly made and fabricated, correctly installed in mining and processing equipment, produce ore and final products reliably and efficiently. Materials engineering, applied during design phases, specification writing, and receipt and installation of materials, ensures that equipment will have the best possible shot at performing up to expectations. Good calls on materials properties - specifying state-of-art materials produced by competent manufacturers - help mine managers to reduce downtime and mining and processing costs. Materials that are improperly made or specified can make a mining operation less efficient or, possibly, uneconomical.

Jan 1, 1992

By Richard M. Crowell, John W. Shield, Todd J. Harvey

Facing lower gold prices, gold mine operators throughout the world are seeking ways to reduce costs for treating refractory sulfide ores. Many gold mines, currently producing oxide ores, have underlying refractory sulfide deposits and the operators could extend the mine life by processing these sulfides. Fortunately, oxidation technologies for refractory ores need not be complex nor costly. GeoBiotics, Inc. has developed heap biooxidation technologies (GEOCOAT™) for treatment of refractory gold ores and concentrates. GEOCOATTM uses simple low cost unit operations common to heap leaching of gold and copper ores. The cost effectiveness and flexibility of GEOCOAT™ heap biooxidation allows mining companies to expand operations to include refract01y ore. In addition, GEOCOAT™ can be utilized to supplement or replace existing oxidation processes such as roasting, autoclaving, or agitated tank biooxidation, and to treat a lower resource grade than is currently economic. The GEOCOAT™ technology involves coating sulfide concentrates onto sized support rocks and biooxidizing the sulfides in a heap configuration. After biooxidation the concentrate is stripped from the support rock, neutralized and leached in conventional cyanidation facilities. Ashanti Goldfields Company Ltd. (AGC) produces over 885,000 ounces of gold per year from oxide and refractory ores at its Obuasi operations in Ghana. Refractory flotation concentrates are treated in roasters and in the world's largest stirred tank biooxidation facility. In August 1998, AGC and GeoBiotics jointly constructed and operated a field demonstration heap of the GEOCOATTM process at the Obuasi operations. The heap contained I 082 tonnes of support rock and 125 tonnes of concentrate. The heap was approximately 19 meters square at the base and 3.8 meters high, and ran for 88 days after which the concentrate was reclaimed. Average gold recovery after biooxidation was 91.5% with a NaCN consumption of 8.6 kg/t. Baseline gold recovery by direct cyanidation of the concentrate was 53.8%.

Jan 1, 1999

By Richard Kolada

If the emission of methane into any particular area of an underground coal mine remains constant, then the calculation of ventilation requirements is relatively simple. This is also true if the ventilation system is recirculatory. However, the emission of methane into a working area can vary considerably. There is always a risk that the ventilation will not be able to dilute the emission of methane to acceptable levels, because there are often short-lived peaks of emission which are very large. The ventilation design must therefore consider what is an acceptable level of risks and calculate the airflow requirements to meet this level. The acceptable level of risk can be calculated by examining the historical variation in methane emission into the area to be ventilated. For example, consider that an acceptable level occurs when methane emission peaks only exceed the capability of the ventilation one day in twenty. Then, from the history of methane emission, it is possible to calculate the necessary ventilation requirements. This method has been used since the early 1980s for conventionally ventilated districts. This paper examines the application of this technique to districts ventilated by recirculatory ventilation.

Jan 1, 1987

By A. Santha Ram

The Indian Mining sector is subjected to a number of trends including call for global transition towards the route to sustainable development. Several regulations were enacted by the Government departments, from time to time, in order to identify the problems faced by the mining industry, including small-scale mining operations and to suggest appropriate mitigation measures in the form of environmental standards and certifications. During this process, a pursuit of sustainable development is regarded as an important area responding to socio-economic conditions, priorities and thus guides eventually corporate policies and practices. Thus, there is a need to formulate strategies for sustainable development. Based on these strategies some environmental indicators were developed in order to track down the progress and to evaluate its success. A case study pertaining to a group of iron ore mines in North Goa and Chromite mines in Sukinda, Orissa was selected for this study. Base line environmental data has been generated in the respective study sites. A core set of environmental indicators monitoring environmental progress based on the P-S-R model pioneered by OECD is linked with indicators for assessing the integration of environmental concerns.

Jan 1, 2003

By Stuart K. Nicol, Malcolm D. Engel, Neville T. Moxon

Ion flotation of gold is a process continuing to be developed to selectively beneficiate dilute aurocyanide liquors using a novel surface­active complexing agent. Aeration of these liquors causes accumulation of the precious metal species in the froth concentrate produced in a flotation column. Successful development of ion flotation for concentrating solids-free liquors would offer a low--cost method for recovering specific metals from heap leach runoff or waste tailings solutions. The cost of constructing and operating a 100 tph gold ion flotation facility has been compared with the relevant carbon adsorption plant costs. Reductions in operating costs of 25% and installed capital of 40% are possible in relevant expenditure areas. For an 'average' processing plant, this equates to annual operating savings of A$ 170 000 and a capital saving of A$ 515 000, with a three-point increase in total (mine and plant) site DCF ROI. High gold inventories are eliminated and the number of processing steps is reduced. The application of gold ion flotation will initially be confined to heap leach operations in the western U.S.A., although use of the process in mine abandonment programmes elsewhere is also an area of future development.

Jan 1, 1992

By Robert Semrad, J. Colin Anderson

INTRODUCTION Sulphur is a mineral which accounts for 0.05% of the earth's crust and may be found either in its elemental state or as a compound. References to sulphur production and consumption are found in early antiquity and international trade of sulphur from volcanic deposits in Sicily began in the 1880's. The supply of sulphur consists of two components: the first is production, and the second is inventory. HISTORICAL SUPPLY Production World sulphur production in 1984 was approximately 53.5 million metric tons, compared to the 1970 level of 42 million metric tons. Of this 11.5 million metric tons production increase, brimstone accounted for close to 90%. Production distribution in 1984 was 62% brimstone, 21% pyrites and 17% other forms, while in 1970 it was 55%, 28% and 17%, respectively. Brimstone Production is derived either from mining operations (Frasch, underground melting and native refined) or as a by-product from hydrocarbons. World production in 1984 exceeded 33 million metric tons, up approximately 10 million metric tons from the 1970 level. This increase was uniformly distributed throughout the world as shown in Table A.

Jan 1, 1986

By C. Newman, N. Schaefer

"INTRODUCTION As underground mining continues to operate at deeper depths and in more complex geologic and geometric conditions, there has been an increased demand for higher design standards in ground control safety and stability within the mining industry. Over the past 20 years, NIOSH has developed, validated, and distributed a number of free software packages (ALPS, ARMPS, AMSS, ARBS, etc.) for the safe design of underground coal mines with respect to ground control. While these programs were initially developed on a, “stand alone model,” where all program data resides on an individual or workgroup computer, modern day web applications have begun to migrate toward a, “multiple-tier model,” allowing multiple users to access an application through an Internet connection. webGroundControl provides users with access to NIOSH ground control design software (ARBS, ALPS, and ARMPS) utilizing a web-based three-tier architecture allowing users to easily access ground control designs, perform on-demand calculations, and provide online collaboration between operational personnel and engineers through a web server. Given advancements in cloud computing as well as the expansion of Internet connectivity on the surface and underground, webGroundControl users will be able to access the applications anywhere (where permissible) at any time. As web-based computing technologies continue to advance and cloud computing gains popularity and momentum, there is an industry need to develop Internet based applications for mine and ground control design (Barczak et al., 2016). The purpose of this paper is to introduce and familiarize those in the mining industry with the basic framework and, more specifically, the user-interface of the webGroundControl application. It is expected that the mining industry will rapidly embrace this new product and webGroundControl will be a new paradigm for mining engineering computer applications. APPLICATION ARCHITECTURE With the advent of cloud computing, web-based applications have become more popular with their adoption of multiple-tier, or n-tier, architecture containing independently operating tiers. The three-tier application architecture utilized in webGroundControl (as outlined in Agioutantis and Newman, 2016) contains a Presentation Tier (Web Server), Logic Tier (Logic/Application Server), and Data Tier (Data Server) as shown in Figure 1. While tiers have the ability to be deployed across and operating on multiple servers, webGroundControl is currently being deployed on a single server located on a university desktop computer. The separation of the application into a series of tiers greatly increases the maintainability and scalability allowing for the adoption of new technologies that can be applied to a single tier without redesigning the software package. The Presentation Tier maintains the user interface translating client requests (through a web server) and displaying results (through a web browser). The Presentation Tier allows for on-demand access from multiple users and compatibility with a variety of personal computers, Smartphones, tablets, etc. on different operating systems. The Presentation Tier has been developed with respect to the HTML and JavaScript languages. Through the utilization of CSS, a coding language which describes the style of an HTML document, the webGroundControl display is able to spontaneously adapt to any browser and any screen. The development of the Presentation Tier will be discussed in more detail in subsequent sections of this paper."

Jan 1, 2017

By Eugene E. Hively

The Tower Kill is a vertical "stirred" ball/pebble mill recently introduced in the Western Hemisphere. This device offers a significant power savings on wet fine grinding applications when compared to conventional tumbling mills. This paper describes the laboratory application of this mill wet grinding hydrated alumina to products of 50 percent passing 2 to 6 microns.

Jan 1, 1983

By E. Poeck, R. Garvey, U. Ozbay, K. Zhang

"A coal bump is characterized by the sudden failure of one or more pillars and an associated release of kinetic energy. Although the geologic conditions surrounding coal bumps are often similar, their occurrence and magnitude are difficult to predict. This paper presents the development of an approach to assess the potential for coal bumps in room and pillar mines through the use of energy concepts and special consideration of the interface properties between the coal and overlying rock. Back analyses were performed on the widespread collapse of a room and pillar mine with an associated 3.9 local magnitude seismic event. Pillar and mine-scale models were constructed using the distinct element method to simulate the mining stages leading up to the collapse. A variety of loading conditions, material properties, and interface properties were evaluated for their effect on unstable failure of single pillars, and the mine-scale model was used to study the evolution of failure during progressive mining. The extent of unstable failure was quantified in these models by calculating the total kinetic energy released during each mining stage. Through the parametric analysis of single pillar models, it was found that softening parameters in either the coal or the coal/rock interface can individually facilitate unstable failure of pillars, but the combination of the two in a single model produced a much higher release of energy during failure. The results of the mine-scale model further illustrated the trend as the combination of softening parameters in the coal and coal/rock interface produced a series of failures which most accurately reflected the collapse event at the mine. The method of room and pillar stability analysis demonstrated in this study effectively explores the potential for large coal bumps. INTRODUCTION The analysis of coal pillar behavior is challenging because natural variability in geologic conditions requires estimation of material strengths and predicted loads for a given mining situation. Large width-to-height (w/h) ratio pillars and deep overburden present an additional challenge, as the failure behavior of such pillars is more complex and may involve unstable, sudden collapse. Empirical methods have improved the success of coal pillar design for a wide range of conditions, but for deep cover and squat pillar geometries, less accuracy has been achieved in predicting pillar performance (Mark 2000). As with any empirical approach, more back analysis and the consideration of new variables may help improve understanding and accuracy of future designs. This paper presents a study of pillar stability through the back analysis of the Crandall Canyon Mine collapse, which occurred in August of 2007. Pillar-scale models are utilized for the calibration of material properties and assessment of potential instability under varying conditions, and mine-scale models are constructed in an attempt to reproduce widespread failure and correlate with observations at the mine. Special consideration is given to the properties of the coal and the coal/rock interface, and kinetic energy is calculated throughout the course of the simulations to gauge the level of instability when pillar failure occurs. Models were constructed using the Universal Distinct Element Code (UDEC) (Itasca, 2014)."

Jan 1, 2016

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Shrinkage stope mining can be employed in steeply dipping veins where both the wall and the vein material are sufficiently strong to stand with only minimal ground support. Shrinkage stope mining takes advantage of the swell in volume produced in breaking in-place ore. This expanded volume of ore is allowed to occupy part of the mined-out vein, and in so doing provides ground sup¬port and a working surface for the miners. Only enough of the ore is drawn after each stoping cut to allow work¬ing room. Not until the entire mining block has been broken is the full amount of broken ore drawn from the stope. DEVELOPMENT PRACTICES FOR SHRINKAGE STOPING The vein is developed into stoping blocks, each of which is 76 to 91 m (250 to 300 ft) in length and 61 m (200 ft) between levels. Level development, described later, is either on the vein or parallel to it with crosscuts periodically driven over to the vein. There are usually two exits from each stope. The first entry is a raise at one end driven either conventionally by raise climber, or by a raised borehole. Present practice is that this raise, put in from level to level, is either a 2.1- x 2.1-m (7- x 7-ft) Alimak raise or a 1.2- to 1.8-m (4- to 6-ft) diam borehole. Either raise system will require pro¬visions for an ore chute carried with the stope to the haulage level or the scram drift. At 9- to 12-m (30- to 40-ft) vertical intervals, dogholes are developed from the raise so that when stoping breaks back into these dogholes, timber or steel sets and necessary service facilities can be installed. The second entry to the stope may be either a similar raise at the stope endline, or more economically, a cribbed raise carried with the stope. As previously stated, development of the ore block is achieved by either drifting on the vein or by paralleling the vein with a footwall lateral and then crosscutting over to the vein at specified intervals. In the following discussion, three basic cases are illustrated in Figs. 1, 2, and 3. In Fig. 1, development is in the footwall by a 3- x 3-m (10- x 10-ft) trackless lateral. A heading is also driven on the ore and taken either full width or driven a standard size and slabbed to the vein width. Approxi¬mately 9-m (30-ft) long crosscuts on 7.6-m (25-ft) cen¬ters are driven from the lateral to the vein. For ease in muck handling they are usually put in on a slant rather than at right angles. A single orepass, located off the lateral and in the center of one stope block, can serve not only that stope, but also one on either side, up to a tramming distance of not more than 137 m (450 ft). Raises are put at the stope endlines on 91-m (300-ft) centers. The back of the vein is shot down and just enough broken material is drawn off to permit access for men to work off the broken ore. Development in Fig. 2 progresses by driving a 3- x 3-m (10- x 10-ft) trackless heading on the footwall side of the vein. The ore block is delineated by raises on 91-m (300-ft) centers. Crosscuts are again driven on 7.6-m (25-ft) centers over to the vein. Approximately four rounds are then taken up to the vein and each round is successively belled out so that all the crosscuts (pockets, in this case) connect. In this development, as opposed to that in Fig. 1 where there were no belled

Jan 1, 1982

By Verya Nasri, Behrouz Esmaeilkhanian, Mohammad Motallebi

The Mount Royal Tunnel (MRT) is a 100-year-old railway tunnel located in Montreal, Canada. The tunnel is approximately 5 km long, 8.7 m wide and 5.5 m high, with a constant 0.6% grade. This existing tunnel will be used by the new Montreal light rail transit network, Réseau Express Métropolitain (REM). This requires the tunnel to be inspected, rehabilitated and made compliant to the recent safety standards. This paper presents the details of the structural inspection and rehabilitation of the tunnel along with the repair methods, addition of a center wall as well as the enlargement of the tunnel section.

Jun 13, 2021

By Stuart M. Jones, Eugene E. Hively

The Tower Mill is a vertical "stirred" ball/pebble mill recently Introduced In the Western Hemisphere This device offers a significant power savings on wet fine grinding applications when compared to conventional tumbling mills This paper describes the laboratory application of this mill In wet grinding hydrated alumina to products of 50 percent passing 2 to 6 microns

Jan 1, 1983

By B. Briggs

Over the past 15 years, the tunneling industry has pushed towards a higher degree of mechanization in nearly every phase of construction. The overall objective of the mechanization was to increase production and the level of safety in tunnel operations. However, the overall complexity of some systems have been a hindrance to a seamless transition to the newer technologies as increased maintenance requirements and re-training of skilled workers has altered the benefit curve for some innovations.

Jan 1, 2010

By Peter J. Szabo

The mining executive works in a cyclical industry. His employment is subject to the vagaries of the marketplace. Boom and bust has always been the industry's story. "Big oil" failed in its mining effort. This left emotional and financial scars on many mining families. One must remember that mining executives always play a high-stakes game. Given the cyclical nature of the mining industry, mining executives need to take care of their own financial security. The mining executive that does not plan for his financial security in this cyclical business is playing Russian Roulette with his future. Take a look at those people who have reached their later years with the mining industry. One usually finds three categories of people. Those few in the first group have been fortunate enough to have worked for one company. They have survived the winnowing process of cyclicality, political problems, and takeovers. They have reached normal retirement age. For them, the retirement years are usually comfortable. People in the second group have worked for those same mining companies. But these people have been forced out by early retirement. Or they have been laid off for cyclical, takeover, or political reasons. The third group includes those who have held many positions in the mining industry. Probably not until their later years are they fully vested, if at all, in a pension program. The first small group, then, is the exception. The others, unless they have taken care of their financial security, are doomed to later years of financial problems, stress, and unhappiness. Too late in life, many realize the harsh realities of failure to plan and implement a personal financial program. Mining executives go to some of the world's leading mining schools, to learn how to make money. The curriculum is technically strong. Most mining graduates are familiar with the intricacies of discounted cash flow, present value, and the Hoskold formula. But few of these same mining schools require courses on personal financial planning. Such courses could give mining students a realistic foundation and the mechanisms with which to cope with their cyclical profession. An executive has only three sources of money. He can make money at work. He can put his money to work. Or he can get money from charities, such as government programs. Of these three sources of money, money at work has the best chance of providing financial security in later years. The mining executive must think of the day when he will be too old to work and contribute to the market place. At this point, the quantity and quality of his investments will determine whether his later years will be secure. Along with the Mining Engineering Handbook, two key books should be included in the mining executive's library. These are Venita Van Caspel's The Power of Money Dynamics and Benjamin J. Stein's Financial Passages. These books contain the basic concepts of personal financial planning so critical to an individual's financial survival. Venita points out the simple formula that is the key to financial independence: Time + Money + American Free Enterprise (Rate of Return) = Opportunity to Become Financially Independent Chart 1 shows that this is not a trite formula. The chart shows the dramatic effects of compounding over time. The chart also shows at various yields what a $2000 a year Individual Retirement Account (IRA) contribution will accumulate. Chart 1 (from Money magazine) is based on the future value of an ordinary annuity. It shows that the key to amassing a small fortune by the time a person reaches his 50s is starting to save methodically in his 20s. This, coupled with the highest yields consistent with safety of principal will ensure financial security. There are significant effects from compounding money over time. Every year the decision is put off to save for later years will cost a person dearly by the time he reaches his 50s. For example, consider the decision to put off for one year a $2000 IRA investment at a 12% return. Starting it at age 31 instead of 30 will cost the executive almost $100,000 by age 65. What if the executive waits until age 35 to get started with his $2000 a year investment? He will then lose nearly $400,000 by age 65. Stated another way, the executive costs himself this $400,000 simply because he did not invest $167 a month for five years. $167 a month - that's less than an inexpensive second car payment. This $400,000 pool of money could provide the executive a $40,000 a year income at age 65, assuming a 10% return on invest-

Jan 11, 1985

By Y. Kurokawa, R. Kurosawa, I. Iwata, I. Fukuoka, N. Sugita

INTRODUCTION The Ningyo-toge Mine is located in the western part of Honshu, about 800 km from Tokyo. Since the discovery of uranium ore deposit at Ningyo-toge in 1955, different kinds of mining and milling tests have been carried out. A new pilot plant of milling and UF6 conversion is also being under construction. Operations are scheduled to start by the end of this year. In 1979, the uranium enrichment pilot plant was constructed at Ningyo-toge and test operations have started. Thus, the Ningyo-toge area is assigned to undertake the Research and Development of the upstream of the nuclear fuel cycle in Japan. PNC has been strengthening its security measures more than ever. This paper will focus on the disposal of wasted material, especially on the disposal of heap leaching residue, in Japan. URANIUM ORE DEPOSITS The uranium of the Ningyo-toge Mine is characterized genetically by epigenetic sedimentary type deposits. They occur in the host rock of Neogene tertiary fluviatile conglomerates and sandstones which unconformably overlies the Mesozoic-Palaeogene granitic basement covered with lucustrine deposits and volcanic ejecta. More than 12 deposits have been found over the area of approximately 160 km2, and distribution of the deposits are strongly controlled by the so-called channel structure, probably remnants of paleo-streams on the denudation surface of the underlying grantic basement. Each ore body has different shape and dimension respectively, but about 2-3 meters thickness with the average grade of 0.05-0.2% U308 is common feature of the ore bodies. The ore is soft and loose, and it is composed of common rock forming minerals of allogenic origin and their altered products. The chemical composition of the ore depends on the constituents of the country rocks or the pebbles. A primary mineral, ningyoite and secondary mineral, autunite are major ore minerals, although three primary tetravalent minerals and thirteen secondary hexavalent minerals have also been detected. MINING EXPERIMENT Since 1959, various types of underground mining tests have been carried out at Ningyo-toge Mine. The room and pillar method was safe and simple, although the uranium ore recovery was below 70%. After experiment of the retreating long-wall method and slicing mining method, it was found that the modified long wall method was the most appropriate for the underground mining of the Ningyo-toge uranium deposits. This method was completed successfully to increase efficiency up to 2 or 3 times, as compared to conventional methods. MILLING EXPERIMENT In 1964, a small test mill was constructed at Ningyo-toge, and a milling experiment using the Ningyo-toge uranium ore was started. Re-modelling of equipments and improvement of the process lead to the establishment of the technique of the so-called wet process, and produce UF4 directly from the ore. The flow diagram is shown in Fig. 1.

Jan 1, 1981

By Charles M. Pelanne

Without insulation we could not survive on this planet. This may sound like a self-serving statement but it is a fact. The blanket of air which surrounds us is a very effective insulating blanket, allowing some shortwave solar radiation to warm us during the day while retaining the longer infrared radiation at night. Our life on earth is very much dependent on materials performing an insulating function without being thought of as insulation, clothing, fur, window glass, the ice of an igloo etc. The mechanisms of heat transfer involved in these materials are analogous to those involved in engineered insulations - Engineered insulations are designed for maximum effectiveness for given conditions within the applicable retraints. Engineered insulations are very effective! As an illustration let us consider this graph (Figure 1). It represents: 1. The radiative heat transfer between two 1 foot2 (0.0932) radiantly black surfaces, one at 100F (38C) and the other at the temperature indicated on the bottom scale. This does not include the convective and conductive heat flow of the air, both of which are significant but small in comparison. 2. The heat transfer resulting from the placement of 1 inch (25.4 mm) of insulation [10 pcf (160 kg/m3) ceramic fiber] between these two plates.

Jan 1, 1976

By G. M. Swinkels

It is generally accepted, that oxygen is a necessary reagent for the flotation of galena with xanthate. Recent measurements have shown (1) that under plant conditions oxygen may be in a short supply to the extent that it limits the rate of flotation. Thus, it becomes important to define the exact role of oxygen and establish a model for the rational design and control of flotation separations, involving oxygen as a reagent. All theories or, the flotation of galena assign oxygen the role of an oxidant. However, they disagree on a number of questions: 1) Is the galena or the xanthate being oxidized? 2) Is a lead dixanthate or dixanthogen the actual collector? 3) Is the oxidation process in which oxygen participates of a chemical or an electrochemical nature? The last question arises from the observation (2) that an anodic bias as well as oxygen will make a galena surface hydrophobic in the presence of xanthate. Theories which are based an the premise that the influence of oxygen is of an electrochemical nature offer elegant models of galena flotation. However, full advantage has not been taken of available electrochemical techniques to discriminate between theories and this is the area towards which this research has been directed (3).

Jan 1, 1970

By P. Zhu

The leaching behavior of copper powders recovered from waste printed circuit boards (WPCBs) was investigated using electrogenerated chlorine and aeration in a sulfuric acid solution with added NaCl and CuSO4. Various parameters were studied to understand the mechanism of the copper leaching. When the current density increased from 0 to 0.1 A/cm2, the temperature from 30 to 60° C, agitation speed from 100 to 400 rpm, and the size ranged from below 75 to 150 µm, the leaching ratios of copper powders increased. The leaching kinetics were analyzed using a shrinking core model: 1-(1-X)1/3 = kt, and the leaching process was chemically controlled with an activation energy of 14.7 kJ/mol due to Cl- ion catalysis. The ratios of [Cu+]/[Cu2+] at various temperatures (30, 40, 50, 60 and 70° C) showed a periodic variation with the increasing of the leaching time. The redox potentials of the leaching solution and copper leaching ratios in different conditions were measured. The result indicates that dissolved oxygen achieved the dissolution of coppers and electrogenerated chlorine can promote the dissolution of copper powders and improve the reaction rate of kinetics in the presence of oxygen.